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Book 3. 



COPYRIGHT DEPOSIT. 



LEAD REFINING BY 
ELECTROLYSIS 



BY 

ANSON GARDNER BETTS 



FIRS T EDITION 
FIRST THOUSAND 



NEW YORK 

JOHN WILEY & SONS 

London: CHAPMAN & HALL, Limited 
1908 



^ 



'{ 



LIBRARY of CONGRESS, 

Two Copies Heceivdd 

FEB 25 1908 

^OowyrigiH £.t«rv. 
<3LAS$>4 XXc. iNu. 
COPY B. 



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Copyright, 1908 



ANSON G. BETTS 



El** &riettttfir #rea* 

Hubert Brummon& attb Cdompamj 

N*tn $ork 



£-6<?£^ 



PREFACE. 

The electrolytic refining of lead bullion has now become 
an established metallurgical process, with further extensions 
confidently expected to come from time to time. Lead 
is almost an ideal metal to refine elect rolytically, because its 
electrochemical equivalent is very high, and hence the power 
cost is small, and the depositing tanks are relatively smaller 
or fewer than for other common metals. Its casting into 
anodes is especially easy, and it stands high enough in the 
electrochemical scale to leave its impurities almost entirely 
in the anode slime, as metals, so there is no appreciable con- 
tamination of the electrolyte. 

The contained information is the result of a number of 
years of study, experiment and practical work, and is pub- 
lished in the hope that it will save those who may be inter- 
ested in lead refining practice or its improvement the re- 
petition of experiments already performed, and give them 
the benefit of the work already done by others and myself. 
Some space has been devoted to theoretical discussions of 
conductivity of electrolyte, etc., which I thought would be 
useful and instructive. 

The variety of methods of slime treatment which are dis- 
cussed in Chapter II, may seem unnecessarily large from the 
practical standpoint, though I myself believe it is desirable 
to treat them at the length I have. I had some hesitancy 



IV PREFACE. 

in including a list of patents published, as they are largely 
my own, but saw clearly that a treatise on this subject re- 
quired all available information of any importance, and would 
be wanted by readers. 

I wish to make grateful acknowledgment to my parents, 
Mr. and Mrs. Edgar K. Betts, of this city, for unfailing assist- 
ance and encouragement while performing my experiments. 
I am indebted for appreciated suggestions and information 
to Dr. E. F. Kern and Dr. Wm. Valentine, who have been 
associated with me in developing process and plant, Dr. Kern 
from April 1902 to June 1904, and Dr. Valentine from Octo- 
ber 1902 until now; to Messrs. W. H. Aldridge, John F. Mil- 
ler, A. J. McNab, and Jules Labarthe of Trail, B. C, Messrs. 
H. A. Prosser, Aug. E. Knorr and Wm. Thum, of the United 
States Metals Refining Co., and Messrs. A. S. Dwight and 
Ernst F. Eurich, and to mam* others. 

Troy, New York. September, 1907 



CONTENTS. 



PAGE 

Preface iii 



CHAPTER I. 

Electrolytes for Lead Refining 

Faraday's law, 3; electromotive forces, 4; rule of electrolytic 
refining, 5; energy requirements, 6 ; f used electrolytes, 7 ; historical, 
Keith's process, 10; Tommasi process, Glaser's experiments, 11; 
development of the Betts process, 12; crystallization prevention, 
14; current efficiency and gelatine, 1G; conductivity of different elec- 
trolytes, 17; acid strength, 18- various solutions, 22; solid lead 
deposition, 22; phenolsulphonate solution, 24; dithionate solution, 
25; preparation of dithionates, 26; fluroborate solution, 28; fluo- 
silicic acid, 29; lead fluosilicate solution, 30; lead fluosilicate, 31; 
dissociation of the solution, 32; losses of fluosilicic acid, 35; silica 
deposited in slime, 36; acid loss on cathodes, 38; acid loss in early 
work, 41; gelatine or glue required, 42; conductivity, 42; metals 
present, 46; Mennicke's experiments, 47; tin in lead bullion, 47; the 
anode slime, 48; polarization of anode slime, 49; e.m.f.'s of solution, 
52 ; limiting current density, 53 ; lead compounds with other metals 
in slime, 54; extraction of lead from very impure bullion, 56; 
Senn's results, 57; iron and zinc, 58; preparation of pure lead, 59. 



CHAPTER II. 

Chemistry of Slime Treatment 60 

Separation by distillation, 60; analyses, 61; amalgamation, 62; 
fusion to alloys, 63 : removal of lead in melting, 64 ; treatment of 
slag, 65; chlorination, 67; chlorination of wet slime, 70; fusion with 
soda, 71; process used at Trail, melting without fluxes, 73; con- 
sideration of electric furnaces for melting, 74; probable power re- 
quired, 75; products of melting, 76; treatment of products, 77; 
melting with sulphur, 78 ; treatment of the slag, electrolysis of slime 

v 



vi CONTENTS. 

PAGE 

as anode, 83; refining slime alloys, 89; wet regeneration process, 
91; fluosilicate solutions, 92; chloride solutions, 92; sulphate solu- 
tions, 93; fluoride solutions, 93; lead peroxide, 93; ferric sulphate 
process, 93; products, 98; treatment of copper slime, 100; electroly- 
sis for regeneration of ferric sulphate, 102; influence of current 
density, temperature, and relative motion of anodes and solution, 
103; deposition of silica on the anodes, 107; diaphragms, 109; 
extraction of the antimony, 111 ; addition of copper to the solution, 
114; perfluoride processes, 115; antimony pentafluoride, 118; use 
of monobasic acids, 119; lead peroxide, 119; use of fluosilicic and 
hydrofluoric acids together, 120; treatment of air-oxidized slime, 
121; alkaline regeneration processes, 123; copper fluosilicate, 125; 
air oxidation of slime suspended in a solution, 126 ; roasting processes, 
128; roasting with sulphuric acid process, 129; dissolving air-dried 
slime in H ? SiF and HF, 134; products of electrolysis, 136. 



CHAPTER III. 

Deposition or Antimony from the Fluoride Solution 138 

Electrolytic refining of antimony, 138 ; deposition from the fluoride 
solutions, using insoluble anodes, 139; anodes, 140; anode reac- 
tions, 141; efficiency, 143; anodes used, 144; impurities, 144; 
analyses, 146; cost of depositing antimony, 148. 



CHAPTER IV. 

Electrolytic Refining of Dore Bullion 149 

Dietzel process, 150; refining with a methyl-sulphate solution, 
152; Moebius and Balbach apparatus, 155; use of gelatine to pro- 
duce solid silver, 158; process used in the Philadelphia mint, 159; 
costs Moebius and Nebel process, 159; plant at Monterey, Mex., 
160; costs, 163; Moebius and Xebel apparatus, 164; attempts to 
deposit solid silver, 166 ; various electrolytes, 166 ; methyl sulphuric 
aeid^ 167; comparative refining costs. 170. 



CHAPTER V. 

The Manufacture of Hydrofluoric and Fluosilicic Acids 174 

Testing fluorspar, 174; small-scale work, 174; retorts, 175; 
condensers, 176; charge, 176; analysis of products, 177; con- 
version to fluosilicic acid, 178. 



CONTENTS. vii 

CHAPTER VI. 

PAGE 

Choice of Constants 180 

Comparison of scries and multiple systems, 180; purity of lead, 
182; cost of glue, L83; current density, 183; tank depreciation, 
IS,"); acid loss, 185; interest on conductors, 186; interest on tanks 
and electrolyte, 180; power cost, 187; final comparison, 189; 
cost of plant, 190; choice of slime process, 191; cost melting with 
sulphur, 192; cost melting to dore, matte, and slag, 193; cost of 
roasting with sulphuric acid process, 194; cost of ferric sulphate 
process, 195. 



CHAPTER VII. 

Refinery Construction, Operation, and Refining Costs 19^ 

Levels in refinery, 197 ; arrangement, 197 ; melting furnaces, 198 ; 
suggested improvement in melting cathodes, 198; dross, 198, 
202; casting anodes, 202; anode mold, 203; closed anode molds, 
209; Trusswell's mold, 209; results in sampling, 213; size of tanks, 
213; concrete tanks, 215; wood tanks, 220; placing of bolts, 221; 
arrangement of tanks, 223; cathodes, 228; casting cathodes, 231; 
cathode-supporting bars, 233 ; foundations for tanks, 233 ; cleaning 
tanks, 234; contacts, 237; circulation of electrolyte, 237; pumps, 
239; electrolyte, 242; washing appliances for electrodes, 244 
washing slime, 246; cranes, 250; floors, 252; evaporators, 252; 
summary of plants, 255; drying slime, 256; melting slime, 256: 
leaching slime, 257; tanks for antimony depositing, 259; ferric- 
sulphate tanks, 260; refinery management, 267; cost of making 
cathodes, 271; cost of tank- room labor, 272; cost of handling lead, 
272; cost of melting lead, 273; cost of refining on a small scale, 
273; comparative costs by the Parkes and Betts processes, 274; 
cost of electrolytic refinery, 279. 



CHAPTER VIII. 

Products 284 

Analyses of bullion refined at Trail, B. C, 284; analyses of Trail 
pig lead, 284; analyses of lead refined by the United States Metals 
Refining Co., 285; silver in pig lead at Trail in early work, 285; 
unequal distribution of silver in cathodes, 286; Trail refined lead, 
286; Trail bullion, 287; slime analyses, 288; products of experi- 
mental refining at Troy, N. Y., 289; lead in Japanese market, 290. 



VUl CONTEXTS. 

CHAPTER IX. 

PAGE 

Treatment of Lead Containing By-products 291 

Refining copper-lead alloys, 291; hard lead, 293; gold-lead 
bullion, 294. 

CHAPTER X. 

Analytical Methods and Experimental Work 295 

Analysis of slime, 295; assay of dore bullion, 297; analysis of 
refined lead, 298; analysis of slag, 302; analysis of electrolyte, 
302; analysis of copper-silver matte, 303; determination of silica 
in slime, 304; analysis of antimony fluoride solution, 304; experi- 
mental work, 305. 

CHAPTER XL 
Bibliography 309 



APPENDICES. 

APPENDIX I. 

Plant of the Consolidated Mining and Smelting Company of 

Canada, Limited, at Trail, British Columbia 312 

Location, 312; power supply, 312; electric machinery, 313; 
subdivision of tank room, 313; tanks and method of lining, 314; 
bus-bars, 315; casting anodes, 316; anode molds, 317; «stacking 
anodes for crane, 318; cleaning scrap, 319; making cathodes, 319; 
melting cathodes, 320; pumping lead. 320; collecting and washing 
slime, 321; report of washing, 322; evaporation of wash-water, 323; 
slime treatment, 323; sodium sulphide extraction, 323; antimony 
depositing, 324; drying and melting leached slime, 325; fluosilicic 
acid plant, 326; labor required, 326; electrolyte, 328; daily report, 
229. 

APPENDIX II. 

Xead Refining Plant of the United States Metals Refining 

Company at Grasselli, Lake County, Indiana 343 

Buildings, 343 ; power plant, 343 ; tank arrangement, 343 ; tanks, 
344; cranes, 344; bus-bars, 345; anodes, 345; melting furnaces, 



CONTENTS. ix 

PAOJ 

345; washing cathodes, 345j melting cathodes, 346; eolUectinc 
ami washing slime. 346 J evaporation of wash-water, 346; electro- 
lyte, 346. 

APPENDIX III. 

Treatment of Lead-refinery Slime with Solution of Ferric 

Fluosilicate and Hydrofluoric Acid 355 

Scale of operation, 355; process used, 356; unsuccessful electroly- 
tic deposition of copper and antimony, 356; use of hydrofluoric 
acid in solution, 357 ; advantages of process, 357; carbon diaphragm, 
357; slime treated, 35S; description of tanks, 360; ferric-iron 
producing tank, 361; solution, 362; results in depositing copper- 
antimony and arsenic, 363; results with ferric-iron tank, 364; 
treatment of slime, 366; extraction of metals, 367, no recovery of 
SiF 6 from slime, 367; improvements in apparatus, 368; metal from 
copper-depositing tanks, 370; direct precipitation of copper from 
the solution, 370; results, 371; precipitation of arsenic and anti- 
mony by lead, 371; products of precipitation by granular lead, 
374. 375; behavior of bismuth in precipitators, 376; separation of 
arsenic and antimony, 377; distinction of products, 378; slime 
treatment, 378; results of slime treatment, 379; influence of HF 
in solution, 380; cotton diaphragms, 381; cathodes, 381; slime 
treatment, 382; proper uae of precipitators, 382. 



LEAD REFINING BY ELECTROLYSIS. 



CHAPTER I. 

ELECTROLYTES FOR LEAD REFINING. 

When two pieces of the same metal are dipped into a 
solution, no difference of electro-motive force is produced 
between the metals, as when dissimilar metals like zinc and 
copper are immersed. When an appropriate solution is used 
and the pieces of metal (electrodes) are placed in an electric 
circuit, metal may be dissolved from one electrode and de- 
posited on the other. The quantities of the various metals 
transported by a certain current in a certain time are pro- 
portional to the atomic weight of the metal, divided by the 
valency in which it exists in the solution (Faraday's law). 
These quantities are, per ampere hour, for a few metals of 
interest to lead refining, as follows. 



TABLE 1. 

Silver 4 . 025 grams per amp. hr 

Lead 3 . 857 

Bismuth 1 . 948 

Antimony 1 . 494 

Copper (Cuprous) 2 . 372 

Copper (Cupric) 1 . 186 

Tin 1 . 105 

Iron 1 . 044 

Gold (Auric) 2.452 



4 . 7 amp. days per lb. 

4.9 " " " " 

9.7 
12.7 

7.95 
15.9 
17.1 
18.1 

7.7 



4 LEAD REFINING BY ELECTROLYSIS. 

As a general thing, by using an appropriate solution, the 
deposited metal (cathode) is pure, although the dissolved 
metal (anode) may be very impure, and on this fact electrolytic 
refining depends. 

Solutions containing a salt of the above metals, generally 
with free acid also present, have been used almost entirely 
as electrolytes. Some of the metals can be got into alkaline 
solution, for example, silver, lead, and copper, and some alka- 
line solutions are used in electroplating, but such solutions 
are not used in refining, so far. 

Only those metals which do not dissolve with evolution 
of hydrogen on immersion in the refining solution have been 
successfully refined up to the present time. For metals which 
cannot be successfully treated wet, as sodium and aluminum, 
fused electrolytes are used. 

The deposition of pure metals depends on the fact that 
each metal has its own electromotive force of solution. The 
electromotive force of solution varies a few hundredths of 
a volt for differences in the concentration of the solution, and 
is somewhat different for different electrolytes. An approxi- 
mation is given in Table 2. This table is practically correct 
for fluosilicate solution. 

TABLE 2. 

Zinc + . 52 volts 

Cadmium + . 16 Cl 

Iron + . 09 " 

Lead -.01 " 

Tin.. -.01 " 

Arsenic — . 40 ft 

Antimony — . 44 " 

Bismuth - .48 

Copper (Cupric) - .52 " 

Silver -.97 " 

Mercury - . 98 " 



ELECTROLYTES FOR LEAD REFINING. 5 

The electromotive force of solution may be denned as that 
difference of voltage which exists between an element and 
the solution also containing the metal in which it is immersed. 
An electric current may be flowing in either direction from 
the electrode and solution, either depositing or dissolving 
metal, without changing the value of this electromotive force 
to any more than a slight extent. 

The results of this are (1) that with an anode containing 
a considerable proportion of that one of the metals present 
in the anode, which stands highest in the series and therefore 
requires the least application of electromotive force to bring- 
it into solution, only that metal will dissolve, and those lower 
in the series will remain in the metallic state, and (2) given 
in the electrolyte a considerable amount of that metal which 
has the lowest electromotive force of those in the solution, 
only that metal will deposit, the electromotive force at the 
cathode being insufficient to deposit the others. 

The rule of electrolytic refining is then, that the metals 
lower in the scale than the principal metal present, are elim- 
inated as metal particles in the anode slime, and the ones higher 
in the series are eliminated as salts dissolved in the solution 
or precipitated from it. 

The elimination as metal in the anode slime is the best 
of the two, as an increasing concentration of other metals 
in the solution requires a change of electrolyte, wdiich is 
troublesome. For instance, in electrolytic silver refining, the 
principal impurity, copper, dissolves from the anode and 
collects in the solution, while the percentage of silver gets 
less. 

Lead, on the other hand, stands higher in the scale than 
all the impurities it contains in appreciable quantities, so that 



b LEAD REFINING BY ELECTROLYSIS. 

the solution does not need to be changed. Taking this into 
consideration, with the great ease of casting lead into anodes, 
and melting cathodes, the comparatively large quantity trans- 
ported by the current, so that a relatively small amount of 
power is necessary and the production is rapid, lead has the 
most favorable physical and electrochemical constants for 
electrolytic refining of all the common metals. 

With an anode of composite metals, we do not have, in 
general, a mixture from which one or more metals may be dis- 
solved, leaving the other metal or metals in the pure state, 
t>ut a mixture of different compounds of the metals between 
themselves. The electromotive force of solution of lead 
combined with antimony for example, is less than that of 
pure lead. The result is that in the electrolytic refining of 
alloys we do not have the full difference in electromotive 
forces of the metals available for making a complete sepa- 
ration. The difference in electromotive force between lead 
and the impurities is though, considerable enough to leave 
something remaining after allowing for the combining force of 
lead and the impurities. The strength of these combina- 
tions varies from practically no combination in the case of 
lead and copper to quite a considerable one in the case of 
lead and antimony. 

The transport of a pure metal from one pure electrode to 
another in the same physical condition, through a solution, 
requires very little energy, provided time is no object. The 
metal of the anode may be, though, in a harder or softer con- 
dition, or may not be the simple metal, but may rather con- 
sist of a series of compounds with other metals present as 
impurities. The elements in these compounds and aggrega- 
tions in general are so weakly united, that usually the energy 



ELECTROLYTES FOB LEAD REFINING. i 

requirement for their decomposition per ton of anode is prac- 
tically negligible. An exception may be noted in the case 
of lead-antimony alloys, "hard lead"; to extract the last of 
the lead from the antimony requires an e.m.f. of over .2 volt. 
The nature of these compounds is of interest as the anode 
slime probably consists of a mixture of them. 

The question of time is, however, one of the most im- 
portant factors, for the refining capacity of a plant of given 
size varies inversely with the speed of working. As we can 
only afford to use a reasonable amount of electric energy per 
ton refined, the first consideration is to find an electrolyte of 
as high an electric conductivity as possible. 

The best conducting electrolyte will be found with a melted 
salt, and melted lead chloride is an exceptionally good con- 
ductor. 

At 580° C, according to Kohlrausch, PbCl 2 has a resist- 
ance of .0373 ohms for a column 1 sq. decimeter by 1 deci- 
meter =.095 ohms per column 1 sq. inch by 1 inch. For com- 
parison, the aqueous electrolyte used with a resistance of 
1.3 — 1.4 ohms is about fourteen times a poorer conductor. 
With the fused electrolyte and the same voltage and separa- 
tion of electrodes, the current density would be about 210 
amperes per square foot, a 4000 ampere vat requiring then 
about 19 sq. ft. of surface. The expenditure of 1-1.5 kilowatt 
would not keep an apparatus of this size, or of one anywhere 
nearly as large, at a red heat, and lj kw. is about all the 
power used for a 4000 ampere tank. It is doubtful if one 
kw. would keep an apparatus occupying more than a few cubic 
inches at the necessary temperature. 

Fused lead chloride dissolves lead sulphide and also gives 
a low melting, high-conductivity electrolyte, which, how- 



8 LEAD REFINING BY ELECTROLYSIS. 

ever, could not be as good as lead chloride alone. Lead 
fluoride I have tried to use as an electrolyte in decomposing 
lead sulphide, but it is relatively infusible. 

Lead chloride melts at a moderate heat, stated in places 
to be about 500° C. Provided a suitable tank could be found, 
if it was attempted to use fused lead chloride with the usual 
depending electrodes, of course they would melt off, and the 
loss of heat would be enormous too. 

Mr. R. H. Sherry made an experiment in my laboratory 
with a mixture of fused zinc and lead chloride, melting below 
the melting-point of lead, so that solid lead electrodes could 
be used. The resistance of zinc chloride is given by Kohl- 
rausch as 10.98 ohms per cubic decimeter, = 27.9 per cubic 
inch. The resistance in Mr. Sherry's experiment was at 310° C. 
about 2.5 ohms per cubic inch, or greater than the aqueous 
electrolytes. 

Special apparatus would have to be devised and the 
current density would have to be far increased beyond 
the 10-15 amperes per square foot used with solutions, to 
reduce the radiating and heat-conducting cross-sectional area 
sufficiently, and this increase of current strength would off- 
set to a greater or less degree, probably greater, the advantage 
of high conductivity. 

Special apparatus has been devised or suggested by Bor- 
chers * and Ashcroft t for refining lead with fused electrolytes. 

The use of a mixture of lead chloride and oxy-chloride 
was proposed by Prof. Borchers,* the idea being that such a 
mixture does not attack iron, while the chloride does. The 



* Electrometallurgy, 1st English Edition, page 338. 

t Electrochem. and Metal Ind., Vol. IV (1906), page 357. 



ELECTROLYTES FOR LEAD REFINING. 9 

crude lead was allowed to flow from groove to groove down 
one side of an iron vessel as anode, with an iron cathode at 
the other side, from which the deposited lead ran down to a 
separate collecting space. Prof. Borchers stated that the 
result in refining lead and bismuth alloys was excellent, which 
I can well believe as far as the chemical result is concerned, 
but that is probably about the only use to which the process 
could be put. After the lead has been largely removed from 
this particular anode metal it is as fusible and liquid as before, 
if not more so, but ordinary crude lead and bullion on the 
other hand would get thicker and less fusible from the accu- 
mulation of copper and arsenic, silver and antimony, and 
would soon be too thick to be handled in this way, long before 
a large part of the lead could be removed. Futher, it would 
take some experimental labor to determine whether all the 
impurities w T ere separated, notably the arsenic and antimony. 
It is also to be much doubted whether the power cost could 
be bought as low as by the wet process. 

Ashcroft has proposed to make the melted lead alloy, con- 
tained in a pot, anode, and spin a cathode of metal above 
the surface of the anode, and very near it. The lead deposited 
on the cathode is to remain suspended by the action of a 
magnetic field, instead of dropping back into the anode metal. 
The magnetic field is to rotate the conducting cathode, which 
it might do, but the action on the lead on the underside of 
the cathode, if there were any action in practice, could not 
act to support this lead, but only to move it in a horizontal 
circle, the same as the cathode itself. 

There will be a difficulty in making the surface of the 
anode metal lie flat, as the metal will tend to move in a circle 
too, from friction and perhaps from magnetism in connection 



10 LEAD REFINING BY ELECTROLYSIS. 

with the current passing through. The same trouble I men- 
tioned before with the impurities of the lead will also appear 
here, to a more serious extent, as the impurities are lighter 
than lead, and as the lead was removed would form a scum 
on its surface. 

The inevitable difficulty with the accumulating impuri- 
ties of the lead in such methods and other serious difficulties, 
made the wet method always seem the best. Since the power 
cost per ton of lead with the cheap electric power now avail- 
able (and this will probably be cheaper as time goes on), is 
only about 50 cents, a great saving is not possible anyway. 

The historical development of electrolytic lead refining, 
up to my own work, is given by Messrs. Watt and Philip in 
their book, " Electroplating and Electrorefining."* 

Prof. N. S. Keith as early as 1878 developed his process 
of refining lead, with an electrolyte containing 180 grams 
sodium acetate per litre, in which was dissolved 18.5 to 22.2 
grams of lead sulphate per litre. He used 20 lb. anodes, 
15X24 inches, and J to ^ inches thick, wrapped in muslin 
cloths to catch the anode slime, which would otherwise drop 
to the bottom of the tank with the refined lead crystals fall- 
ing from the cathodes. 

At Rome, N. Y., a plant was built with 30 tanks produc- 
ing 3 tons of lead per day of twenty-four hours. The tanks 
were circular, made of a kind of concrete mixture, 6 feet in 
diameter, 40 inches deep, with a central pillar 2 feet in diam- 
eter occupying the centre of the tank. Brass cylindrical 
cathodes were used 2 inches apart, and extended all the way 
round the tank, with 270 anode plates to the tank 6X24 

* New York and London, 1902. 



ELECTROLYTES FOR LEAD REFINING. 11 

inches, and weighing 8 lbs.; current was supplied by an Edison 
dynamo of 2000 amperes and 10 volts. The anodes were 
hung from a frame which rotated continuously and carried 
scrapers that scraped the deposited lead from the cathodes. 
The current density, calculated from these figures, was 3.2 
amperes per square foot. 

Tommasi * published various articles in 1897 and 1898 
describing his arrangement for refining lead, also with the 
acetate solution. His proposition was to use as a cathode a 
circular aluminum-bronze disc, mounted on a shaft just above 
the top of the electrolytic cell, which disc was to turn once 
a minute, and be relieved of its deposit of spongy lead by a 
scraper above the tank, while the spongy lead was automati- 
cally carried off to a press. Tommasi, in elaborate but wrong 
calculations, presumes a refining cost of 8.6 francs per metric 
ton with steam power and 5.8 francs with water power. The 
process described would, however, probably cost nearer 50 
or 75 francs, if all went well. 

L. Glaser f reports a number of experiments with little 
exactness of description, in depositing lead from various 
electrolytes, the description being limited to lead' nitrate, 
lead nitrate and sodium nitrate, lead acetate, sodium nitrate 
saturated with lead hydrate, and caustic potash with lead 
hydroxide in solution of various strengths, and claims a solid 
lead deposition. I have repeated Glaser's experiments very 
fully as far as it is possible to follow him, and in no case 
was I able to get a solid deposit of any measurable thickness. 



* Comptes Rendus, 1896, Vol. 122, p. 1476. Zeitschrift fur Electrochemie, 
Vol. 3, 92, 310, 341. 

t Zeitschrift fur Electrochemie. 1900, Vol. 7 (24), 365-369 and (26) 
381-386. 



12 LEAD REFINING BY ELECTROLYSIS. 

Following a work of Foerster and Guenther, who offered 
the explanation that spongy zinc deposits are caused by the 
simultaneous deposition of zinc oxide with the zinc, Glaser 
attempts to prove the cause of the loose lead deposit to be 
due to the co-deposition of lead hydroxide. This is, how- 
ever, incorrect theory, as anyone can easily see by electro- 
lyzing lead solutions containing free acid,- as nitric, acetic, 
fluosilicic, etc., which by their acidity absolutely prevent the 
separation of lead hydroxide, and yet give loose deposits. 
It is also possible, without making any alteration of the 
acidity of a proper solution, to cause the separation of a 
solid instead of incompact deposit, as will be seen later. 

The next proposition for refining lead is seen in patent 
specifications.* I refined about half a ton of lead, in 4 cells 
each 10J" wide, 16" deep, and 30" long, containing 9 anodes 
weighing about 12 lbs. each, and 10J inches wide by 13J 
inches deep. The strength of the solutions varied from 4 
to 20 grams lead and 12 to 25 grams SiF 6 per 100 cc, but the 
deposit was always incompact. The cathodes consisted of 
sheet iron, which it was attempted to coat with lead by 
dipping into lead in a deep pot, and afterward by lead- 
plating them. In the first experiments the idea was to 
simply melt the lead off the iron when the cathodes were 
finished, by dipping into melted lead, after which the cathodes 
could be returned to the tanks. In the later experiments 
the cathodes were greased and the lead afterward peeled off 
mechanically. 

Every few hours during the runs, which lasted during the 



*U. S. Patents, A. G. Betts; 679,357, July 30, 1901; 679,824 August 6, 
1901. 



ELECTROLYTES FOR LEAD REFINING. 13 

day time for about a week, with a current from 120-150 
amperes ( = 7 to 8.8 amperes per square foot, total e.m.f. per 
cell 0.175 volts), the cathodes were taken out and passed 
through steel rolls of about 3" diameter. The sheets came 
through the rolls in quite a solid deposit and with a smooth 
surface. A good deal of electrolyte was squeezed out and 
part of this was lost, and the whole was a disagreeable job 
witrT the machinery at hand. A sample of the deposit, 
which seemed quite solid, showed a specific gravity of 10.28 
only, against 11.36 to 11.40 for pure lead. This would 
mean a loss of electrolyte in the remaining pores, per ton 
refined, of about .3 cubic foot, still a rather serious item. 

The idea was to equip the tanks, as may be seen from 
the above-mentioned patents, with a pair of rails on each 
side, over which ran a machine that automatically stopped 
over each cathode in succession, raised it through a pair of 
rolls and returned it to its position in the tank. 

TABLE 3. 
Analyses of Bullion Treated and Refined Lead Produced. 

Bullion. Refined Lead. Slime. 

Ag about .50% Ag .0003% Ag 36.4% 

Cu " .31% Cu .0007% Cu 25.1% 

Sb " .43% Sb .0019%, Sb 29.5% 

Pb " 98.76% Pb 99.9971% Pb 9.0% 



Bullion. Refined Lead. 

Cu 75 % Cu .0027% 

Bi 1 .22 % Bi .0037% 

As 936% As .0025%, 

Sb 6832% Sb .0000% 

Ag 358.89 oz. Ag .0010% 

Au 1 . 71 oz. Au None 

Fe .0022% 

Zn .0018% 



14 LEAD REFINING BY ELECTROLYSIS. 

The idea of using rails at the side of the tanks, over 
which carriages may be run to carry electrodes in and out 
and slime out, seems to be one that might be adopted in 
refineries with some advantage. 

The objection to a loose mass of separate lead crystals, 
as previously invariably produced in electrolyzing lead solu- 
tions, is serious from the refining standpoint. After doing 
considerable work with mechanical methods of compacting 
the lead, I discovered certain materials that, if added to such 
a solution as the fluosilicate, caused the production of solid 
lead deposits, notably gelatine and pyrogallol, although when 
added to acetate solutions they had no valuable effect. As 
gelatine is the cheapest, it alone has been adopted in prac- 
tice. Saligenin and resorcin were found to cause an im- 
provement, but not quite so solid a deposition as the other 
two. The search was not limited to organic reagents, but 
they alone were found suitable. With the addition of small 
amounts of gelatine to a fluosilicate solution, perhaps 1 part 
of gelatine to 5000 or less parts of solution, the lead sepa- 
rates as a solid smooth deposit, with a specific gravity of 
11.3 to 11.4, the same as cast lead. 

The way in which the gelatine, etc., bring about this re- 
markakle result is hard to trace. I am satisfied that the 
next step toward a complete explanation is to be found in 
variation in hardness or tensile strength of the cathode deposit 
resulting from the use of gelatine, etc. The principal rea- 
sons for this opinion are based on these facts: 

1. * Although equally pure, the solid electrolytic lead 
deposit is several times stronger than ordinary lead. 

* Betts, Trans. Am. Electrochem. Soc. Vol. VIII, 1905, page 83. < 



ELECTROLYTES FOB LEAD REFINING. 15 

2. The greater the tension at the surface of an electro 
deposit, the greater the tendency to keep the new surface 
forming smooth.* 

3. Lead deposited from liquids in which its surface ten- 
sion after immersion must be greater, is smoother, e.g., pyri- 
dine solutions, f 

4. Strong metals, otherwise suitable for electro-deposition, 

give the smoothest deposits. Weak metals give loose crys- 
talline growths. 

In what way the gelatine or other similar addition acts 
to increase the strength of the surface layer of cathode 
deposit has not been definitely established. 

It is an interesting fact that the addition of gelatine or 
pyrogallol to the acetate and similar solutions does not 
cause the production of a solid deposit, while the addition 
of gelatine in the strong-acid solutions, fluosilicic, fluoboric, 
etc., does. 

Snowclon claims to mechanically produce solid lead from 
the acetate solution by the use of a rapidly revolving cathode, 
but does not give the specific gravity of the product. | 

I criticize the practice of describing lead deposits as 
solid, homogeneous, etc., without making any definite state- 
ments as to the specific gravity, mechanical soundness, etc. 
Some definite standard is required to show how " solid" a 
deposit is, also the thickness of the deposit should be detailed. 
Many deposits of slight thickness have quite a smooth and 
solid appearance for that reason, but after building them up 
a little more, their true loose nature can be recognized. 



* Betts, Trans. Am. Electrochem. Soc. Vol. VIII, 1905, page 85. 

f Kahlenberg, Trans. Am. Electrochem. Soc. Vol. VI, 1904, page 40. 

J Trans. Am. Electrochem. Soc. Vol. IX, 1906, page 221. 



16 LEAD REFINING BY ELECTROLYSIS. 

The lead deposit forming in lead fhiosilicate-fluosilicic 
acid solutions, containing .1% of gelatine, and five or more 
per cent lead, is smooth and solid, and thick pieces cut from 
the deposit show a specific gravity of 11.35 to 11.40; the 
same metal after melting and casting shows practically the 
same specific gravity, in some cases exactly the same. With 
a little more lead, say 7-8%, and the average current density 
employed in commercial operations of 15 amperes per square 
foot, the resulting cathodes, after reaching a considerable 
thickness, are smoother. The electrochemical equivalent of 
lead is so high that with only 5% lead, the layer in the imme- 
diate neighborhood of the cathode is probably nearly ex- 
hausted in respect to lead, and if the lead is allowed to go 
much below 4%, a black, slimy deposit of lead is the result. 

That gelatine does not affect the current efficiency I 
determined some years ago in the following way. Two solu- 
tions were electrolyzed in series, one containing gelatine and 
the other without. 

TABLE 4. 



Experiment 
No. 



Without Gelatine. "Uith Gelatine. 



Weight Weight Weight * Weight 

Deposited. Dissolved. Deposited Dissolved. 



1 


4, 


06 


gr. 


4 


. 0* 


gr. 


4 


04 


gr. 


4 


.06 


gr- 


2 


24. 


70 


gr. 


24 


.89 


gr- 


24. 


72 


gr. 


24 


.90 


gr. 



The electrodes were arranged to be weighed without re- 
moving them from the solution at all. to avoid the disturb- 
ing influence of air oxidation on the spongy deposit from the 
solution with no gelatine. 

The amount of lead transported by the current has been 



ELECTROLYTES FOR LEAD REFINING. 17 

made a careful study.* Under the most perfect conditions 
yet applied to the deposition of lead, the electrochemical 
equivalent of lead is found to be 103.43, that is, the atomic 
weight of lead corresponding is 206.86 and the amount of 
lead transported per ampere hour is 3.857 grams, f 

For refining lead, we require a solution of as high a 
conductivity as is commercially available, which also will 
contain at least several percent of combined lead without 
being saturated with the lead salt. Certain acids, such as 
fluosilicic, fluoboric, dithionic, various fatty sulphuric acids, 
as ethyl-sulphuric acid, and phenol-sulphonic and benzene 
sulphonic acids, have been found to meet the requirements 
of high electric conductivity and solubility of their lead salts. 

For a comparison of the conductivity of these acids with 
the acetate electrolytes of Keith and Tommasi, see Table 5. 

TABLE 5. 

Approximate 
In 100 c.c. Solution. Name. Temperature. Resistance 

per Inch Unit. 

7.7% Pb (C 2 H 3 2 ) 2 Acetate 19.6° C. 75 ohms. 

$14.5% Pb (C 2 H 3 2 ) 2 Acetate 19.4° C. 58 

5 gr. Pb, 7 gr. BF 4 Fluoborate 25 ° C. 4 

5 gr. Pb, 15.7 gr. C 6 H 5 S0 3 Benzenesulphonate. .25 ° C. 2.7 " 

5 gr. Pb, 12.5 gr. C 2 H 5 S0 4 Ethylsulphate 25 ° C. 3.6 " 

5 gr. Pb, 9.5 gr. C 2 H 3 2 Acetate 25 ° C. 84 

15.7 gr. Pb, 2.4 gr. K, and 

21.4 gr. C 2 H 3 2 Acetate 26 ° C. 22 " 

Considerations of cost so far have required the use of 
fluosilicic acid, but dithionic acid may yet be found to be 
more economical. 

Several recent writers have apparently thought that 
fluosilicic acid had some peculiar property that made it better 

* Betts and Kern, Trans. Am. Electrochem. Soc. Vol. IV, 1904, page 67. 
f F. W. Clarke, Trans. Am. Chem. Soc. Vol. XXVIII, 1906, page 307. 
X Kalender fur Elektrochemiker, 1903. Neuberger. 



18 LEAD REFINING BY ELECTROLYSIS. 

than other acids for giving a solid lead deposit. Mr. Serin* 
in his paper describes experiments to see if it was also suitable 
for refining cadmium, and obtained some excellent results 
with fluosilicate of cadmium, and H. Mennicke f has applied 
the acid to refining tin and obtained good results. These 
views are, however, probably incorrect. The true suitability 
of an acid is more dependent on its strength, and solubility 
of its salts, than on other things. 

Messrs. Senn and Mennicke would probably have done 
equally well with other non-oxidizing acids having an equal 
strength and forming salts of cadmium and tin respectively, 
of equal solubility. 

Prof. Ostwalcl in his work " Outlines of General Chem- 
istry," translated by Dr. James Walker, New York, 1890, 
page 360, gives a valuable table and makes this statement 
in connection with it. 

" The fact stated by Hittorf that the power of reaction 
and the electrolytic conductivity are always concurrent 
properties, speaks at once in favor of this assumption. [That 
the strength of an acid is a definite and definable character- 
istic proportional to its dissociation.] It obtains further 
support from the circumstance that the processes of electro- 
lytic conductivity and of chemical decomposition both de- 
pend on the molecules under consideration falling into 
smaller sub-molecules; without this decomposition there can 
neither be a new distribution of parts as in chemical reaction, 
nor a transport of electricity attached to the ions, as in con- 
duction. 



* Zeitschrift fur Electrochemie. II (1905), 229-245. 

t Zeitschrift fur Electrochemie. XII (1905), 112, 136, 161, 180. 



ELECTROLYTES FOR LEAD REFINING. 19 

" But the most decisive and telling argument for the 
soundness of the assumption is the numerical agreement of 
the values for the chemical activity on the one hand and the 
electric conductivity on the other. The numbers on pages 
354 and 356 for the rate of catalysis of methyl acetate and 
of the inversion of cane sugar, agree so closely with those 
representing the relative electric conductivity, that there 
cannot exist the slightest doubt of the intimate connection 
between the two series. 

" In the following table there is tabulated under I. the 
electric conductivity of normal solutions of acids, under II. 
the coefficients of velocity for the catalysis of methyl acetate, 
and under III. the coefficients of inversion of cane sugar. 

TABLE 6. 

Acid. I. 

1. Hydrochloric, HC1 100 

2. Hydrobromic, HBr 100. 1 

3. Nitric, HN0 3 99.6 

4. Ethanesulphonic, C 2 H 5 S0 2 OH 79.9 

5. Isethioniq, C^OH -S0 2 OH 77.8 

6. Benzenesulphonic, C 6 H 5 -S0 2 OH. . 74.8 

7. Sulphuric, H^O* 65. 1 

8. Formic, H -COOH 1 . 68 

9. Acetic, CH 3 -COOH 424 

10. Monochloracetic, CH 2 C1 COOH. . . 4.90 

11. Dichloracetic, CH1 2 COOH 25.3 

12. Trichloracetic, CC1 3 COOH 62 . 3 

18. Lactic, C^OH -COOH 1 . 04 

25. Oxalic, (COOH) 2 19.7 

29. Tartaric, C 2 H 2 (OH)(COOH) 2 2.28 

32. Citric, C 3 H 4 OH(COOH) 3 1 .66 

33. Phosphoric, PO(OH) 3 7.27 

34. Arsenic, AsO(OH) 3 5 . 38 



II. 


III. 


100 


100 


98 


111 


92 


100 


98 


91 


92* 


92 


99 


104 


73. 9f 


73.2 


1.31 


1.53 


.345 


.400 


4.30 


4.84 


23.0 


27.1 


68.2 


75.4 


.902 


1.07 


17.6 


18.6 


2.30 




1.63 


1.73 




6.21 

4.81 





* Should be .98 , 

,„, . , , ,, _ ason page 3o4 of Ostwald's book. 

f Should be o4.7 J r & 



20 



LEAD REFINING BY ELECTROLYSIS. 



We are especially interested in the strength of fluosilicic 
acid, fluoboric acid, and dithionic acid, as well as some of 
those given in the table. 

Mr. R. H. Sherry made determinations of the strength 
of these by the methyl-acetate method as described in Ost- 
wald's same book, page 352, and also made tests on HC1 and 
H 2 S0 4 as a check. His results cannot be directly added to 
Ostwald's table, as they were made at different temperatures 
and a different amount of methyl-acetate was used. 

He used normal solutions of H 2 SiF 6 , that is, containing 
7.2 gr. H 2 SiF 6 per 100 cc; normal solution of dithionic 
acid, 8.3 grams H 2 S 2 6 per 100 cc; normal sulphuric acid 
4.9 grams per 100 cc; normal hydrochloric acid 3.65 grams 
HC1 per 100 cc Through an error If N fluoboric acid was 
used instead of normal, and the only basis of comparison 
made was with 1§N HC1. Normal fluoboric acid = 8.8 grams 
BHF 4 per 100 cc 

The figures give the amount of acetic acid liberated in 
grams in 90 minutes and are very nearly proportional to 
the strength of the acids. 

TABLE 7. 



At Approximately 26° C. 


At 26.5- 


-27° C. 


N H2SO4 


N H 2 SiF 6 


NHC1 


N H2S2O6 


1§N HC1 


lfN BHF4 


(0) .2330 
(b) .2387 


(a) .2540 ; (a) .4106 

(b) .2578 (6) .4116 


(a) .4223 
(6) .4169 


(a) .5541 

(b) . 5568 


(a) .5450 

(b) .5391 



We then get approximately the ratios of the following 
table, taking normal HC1 = 100: 



ELECTROLYTES FOR LEAD REFINING. 21 



TABLE 8. 

1. Hydrochloric acid, HC1 100 

2. Dithionic acid, H^A 102 

3. Fluboric acid, BHF 4 95 

4. Fluosilicic acid, H^iFg 62 

5. Sulphuric acid, H;>S0 4 57 

6. Acetic acid, HCH~C0 2 345 

7. Ethyl sulphuric, C 2 H 5 S0 4 H 74 

8. Benzene sulphonic, C 6 H 3 S0 3 H 74 



In other tables in his book Professor Ostwald gives the 
strength of benzene-sulphonic acid and ethyl sulphuric acid, 
as determined by the methyl acetate method, as practically 
100. The figures in this table are for determinations made 
by the electric conductivity method. I do not think the 
methyl acetate method is reliable for acids having an organic 
residue on account of the naturally greater dissolving power 
such acids must possess even in solution, for organic sub- 
stances as methyl acetate. Such determinations certainly 
do not check anyway with the conductivity determina- 
tions. 

In recent experiments, all these strong acids have been made 
up into lead-depositing electrolytes containing 4 grams and 
more of lead per 100 cc. beside free acid, giving lead deposits 
of varying characteristics, but all of them always loose and 
crystalline and unsuitable for practical work, on account of 
their lack of solidity and the short circuits produced. All 
the electrolytes in the following list and many others, too, 
have produced loose deposits without exception. 

A partial list of electrolytes used for depositing lead is 
given in Table 9, the figures being for grams per litre. 



22 LEAD REFINING BY ELECTROLYSIS. 

TABLE 9. 

400 grs. lead nitrate 

60 grs. lead nitrate, 33 grs. sodium nitrate, 6 grs. nitric acid 

250 grs. lead nitrate, 35 grs. sodium nitrate, 6 grs. nitric acid 

350 grs. lead nitrate, 35 grs. sodium nitrate, 6 grs. nitric acid 

33 grs. lead nitrate, 33 grs. sodium nitrate, 6 grs. nitric acid 

33 grs. lead nitrate, 100 grs. sodium nitrate, 6 grs. nitric acid 

100 grs. lead nitrate, 400 grs. sodium nitrate, 6 grs. nitric acid 

530 grs. lead acetate, 117 grs. ammonium acetate, 33 grs. acetic acid. 

' 400 grs. sodium nitrate satuarted with lead hydrate 

5.6 grs. caustic potash saturated with lead hydrate 

448 grs. caustic potash saturated with lead hydrate 

50 grs. lead, 87 grs., BF 4 

50 grs. lead, 157 grs. benzene sulphonic acid radicle 

50 grs. lead, 125 grs. ethyl sulphuric acid radicle 

27 grs. lead formate, 46 grs. formic acid 

60 grs. lead acetate, 60 grs. acetic acid 

32.5 grs. lead acetate, 60 grs. acetic acid, 60 grs. potassium acetate. 

186 grs. lead lactate, 93 grs. lactic acid 

186 grs. lead lactate, 186 grs. lactic acid 

With the addition of gelatine to the strong acid solu- 
tions (fluosilicic, fluoboric, dithionic, organic sulphuric and 
sulphonic acids), they give solid lead deposits, the best of 
which have been obtained with fluosilicic and fluoboric acids, 
and on one occasion, when the solution happened to be in 
just the right condition, with dithionic acid. Benzene-sul- 
phonic acid gives the roughest deposits and is the most 
troublesome to use. The lead salt is not very soluble. 

A number of these deposits of considerable thickness have 
been examined for specific gravity from time to time, with 
these results: 

TABLE 10. 

Fluosilicate 11 .29 11 .35 11.36 

Benzenesulphonate 11.35 . 11.37 

Ethyl sulphate 11.27 11.31 

Fluoborate 11 . 39 

Dithionate 11 . 20 

Phenolsulphonate 11 . 35 



ELECTROLYTES FOR LEAD REFINING. 23 

Phcnol-sul phonic acid gives an excellent lead deposit, as 
does benzene disulphonic acid and phenol disulphonic acid. 
All the other organic sulphonic acids that I tried, as toluene 
and naphthalene sulphonic acids, give altogether too insolu- 
ble lead salts. Methyl- and amyl-sulphuric acid are prac- 
tically equivalent to ethyl-sulphuric acid. Ethane disulphonic 
acid, from C 2 H 4 Br 2 and ammonium sulphite by Stacker's 
reaction, gave too insoluble a lead salt. COCl 2 and a sul- 
phite solution did not give the expected dioxy methylene 
disulphonic acid. Calcium carbide and concentrated sul- 
phuric acid gives a number of sulphonic acids, but I have 
not investigated this reaction to a great extent. 

With tax-free alcohol there is a slight chance of economic- 
ally using ethyl-sulphuric acid. The reaction between alcohol 
and sulphuric acid is 

C 2 H 5 OH + H 2 S0 4 = C 2 H 5 S0 4 H + H 2 

and provides a relatively cheap acid for refining lead. Ethyl- 
sulphuric acid in strong solution decomposes, however, again 
into alcohol and sulphuric acid. I accordingly determined 
the decomposition rate of a solution I had, which contained 
10 grams lead and 8.8 grams free C 2 H 5 S0 4 H per 100 cc. 
This solution deposited about .03 grams lead sulphate per 
day, at about 25° C. This corresponds to about 2.1 lbs. 
C 2 H 5 S0 4 H decomposed per ton lead refined. To prepare 
2.1 lbs. of the acid would require about 1 lb. alcohol and 2 
lbs. fuming H 2 S0 4 (30% S0 3 ). The alcohol would cost at 
least 4 cents and the sulphuric acid 2.5 cents, or a total 
cost per ton lead for these materials of 6.5 to, say, 10 cents. 
The above solution was pretty weak, however, and the tern- 



24 LEAD REFINING BY ELECTROLYSIS. 

perature a little low, so I think it would be found with such 
a solution as would be used practically, that the decompo- 
sition would be three or more times as great. The resistance 
of this solution of lead ethyl sulphate and ethyl sulphuric 
acid (10 grams lead and 8.8 grams C2H5SO4H per 100 cc.) 
was at 27° C, 2.6 ohms per cubic inch as against about 1.4 
ohms for the regular fluosilicate solutions. 

Solutions of lead phenol-sulphonate gave excellent re- 
sults as far as conductivity and solid lead deposition went, 
but the solution seemed to be unstable for a crystalline 
deposit kept forming for a long time. The sulphonation of 
"the phenol takes place readily and good yields may be ob- 
tained. 20 grams of phenol were heated up to 180° C. for 
one hour with varying quantities of H 2 S0 4 . With 25 grams 
H2SO4, titration of the product with sodium carbonate, 
showed that the reaction was nearly quantitative. With 
more sulphuric acid, considerable disulphonic acid was 
obtained, about 30% of the monosulphonic acid being con- 
verted to disulpho-acid, when 40 grams H 2 S0 4 were used for 
20 grams phenol. 

A solution containing 75% mono-acid and 25% di-acid 
and 30 grams lead in 100 cc, of the following composition: 
lead 30 gr., C 6 H 5 S0 3 ' 28.4 gr., and C 6 H 4 (S0 3 )2" 15.2 gr. per 
100 cc. gave a resistance per cubic inch of 2.04 ohms. 

The solution was practically neutral and the resistance 
would no doubt have been much less with more free acid. 
However, the relatively high cost of pure phenol, and the 
difficulty I found in trying to get a suitable solution from 
crude phenol or cresol, led to the abandonment of these ex- 
periments, although they looked promising at first. 

Lead benzene-sulphonate is relatively little soluble and 



ELECTROLYTES FOR LEAD REFINING. 25 

the lead deposit was poor. It is also more difficult to get 
even a fair yield of benzene sulphonic acid. 

Many tests have been made with dithionic acid electro- 
lytes, and on one occasion a very excellent deposit was got. 
All the other experiments have given a rather poor deposit. 
The surpassing conductivity of dithionic acid, the fact that 
the only raw material actually necessary to make it is SO 2 
which so many works have plenty of and to spare, make it 
seem almost an ideal electrolyte. The acid is subject to 
decomposition, however, in strong or warm solution, as fol- 
lows: 

H 2 S 2 6 = H 2 S0 4 + S0 2 , 

both the products of reaction being bad. The sulphuric acid 
precipitates lead sulphate into slime, but worst of all, the S0 2 
is reduced by the cathode, forming lead sulphide. 

S0 2 + 3Pb + 2H 2 S 2 6 = 2PbS 2 6 + 2H 2 + PbS, 

which spoils the cathode deposit, if deposited in any quan- 
tity. The one good deposit I mentioned probably resulted 
from the use of a solution freshly made up with crystallized 
lead dithionate, water, and dilute sulphuric acid to precipi- 
tate out part of the lead and set free some of the dithionic 
acid, which contained no S0 2 . 

I had experiments made lasting for several week's con- 
tinuous run with another solution, but the deposit was " door- 
mat" to the last, but I am not satisfied that the solution, if 
used properly, cannot be made to yield an excellent deposit.* 



* Since the above was written further experiments have also failed to 
give an entirely satisfactory deposit continuously. 



26 LEAD REFINING BY ELECTROLYSIS. 

The rate of decomposition is quite slow. A solution con- 
taining 6.6 grams Pb and 5.75 grams free H 2 S 2 6 , or a total 
of 10.75 grams S 2 6 per 100 cc, giving a resistance per 
cubic inch at 26.5° C. of 1.92 ohms (a nuosilicate solution 
of corresponding, acidity would be about 2.6-2.7 ohms), de- 
composes at the rate of complete decomposition of the S 2 6 
in about 80 weeks, or from 1.75 to 2.2 lbs. H 2 S 2 6 decom- 
posed per ton of lead refined. Another solution containing 
7.5 grams lead and 14.4 grams S 2 06 per 100 cc, decomposed 
in 5J months at the rate of total decomposition in 36 years. 
The conductivity of a solution containing 7.5 grams Pb and 
12.6 grams S 2 6 " per 100 cc. at 21}° was 1.75 ohms. 

The preparation of the dithionic acid we used was accom- 
plished in two different ways. In each case Mn0 2 was dis- 
solved in water by a current of S0 2 gas passed through. Two 
reactions may take place as follows, of which the first is the 
only useful one: 

Mn0 2 + 2S0 2 = MnS 2 6 , 
Mn0 2 + S0 2 = MnS0 4 . 

Conditions favoring the first are low temperature, say 
10° C, and the continual presence in the solution of an excess 
of S0 2 . Under these favorable conditions the yield has been 
as high as 86% of the manganese dissolved converted to 
dithionate and 14% to sulphate, or a yield of 93% on the S0 2 
used. In another case the yield on sulphur was 81.6%, and 
on manganese dissolved 63.3%. The reaction between the 
manganese dioxide and H 2 S0 3 is rapid. 

At first the manganese salt was decomposed with lead 
peroxide: MnS 2 6 + Pb0 2 = Mn0 2 + PbS 2 6 . It was soon 



ELECTROLYTES FOR LEAD REFINING. 27 

found that while this reaction was all right when sulphates 
were absent from the solution, the yield was poor otherwise. 
We accordingly precipitated the S0 4 first by adding lead 
dithionate from a previous batch. It was thought that the 
manganese dioxide precipitate could be used over and over 
again, but the precipitated dioxide will not give nearly as 
good a yield as native pyrolusite. 

Only certain varieties of lead peroxide will react with the 
lead solution. The peroxide precipitated by heating a mixed 
solution of lead acetate and calcium hypochlorite, did very 
well, but the cost would be too high for practical work, so 
I devised an electrolytic method as follows: By electrolyzing 
a solution of common salt with carbon cathode and lead 
anode, lead hydrate is precipitated, especially well if the solu- 
tion is heated a little. If the mixture is then electrolyzed 
with carbon anode and lead cathode, sodium hypochlorite 
is produced, which converts the lead hydrate into lead peroxide. 
The original idea was to merely reverse the current occa- 
sionally, but that did not do very well because there was 
always a coating on the lead anode, and when the current 
was reversed the coating was again reduced to spongy lead 
with a corresponding loss of efficiency. The difficulty was 
surmounted by using two sets of electrodes in different parts 
of the cell, and even then there was difficulty with the for- 
mation of a coating on one of the electrodes, but an obser- 
vation of Dr. Kern that the coatings fell off if the current 
was interrupted entirely for a short time occasionally, put 
us in possession of a practicable process of preparing the pre- 
cipitated lead peroxide from lead anodes by the help of elec- 
tricity. The product was entirely free from Pb(OH) 2 if the 
proportion of the two sets of reactions we were carrying on 



28 LEAD REFINING BY ELECTROLYSIS. 

were so adjusted that there was always excess of NaOCl 
formed. 

The other method of preparation is based on treating the 
manganese dithionate and sulphate solution with slacked 
lime, giving a solution of the calcium salt, and a precipitate 
containing the manganese, which could perhaps be used 
equally as well or better than the original Mn0 2 ore used, 
in a Spiegeleisen or ferromanganese furnace, thus paying for 
the manganese. The calcium salt was decomposed with sul- 
phuric acid for calcium sulphate and dithionic acid. I had 
experiments made in my laboratory on this process, but the 
results do not show anything for or against its probable 
success. 

Fluoboric acid is a somewhat better conductor than fluo- 
silicic acid, if the comparison is made on the basis of equal 
neutralizing power, in about the ratio 3 to 2 for weak solu- 
tions, the difference becoming less as the solutions become 
stronger. The amount of HF required to produce the acids 
in the ratio for equal acidity, is 80 to 61. For weak solutions 
for a given amount of fluorine, a slightly greater conduc- 
tivity can be secured by the use of boric instead of silicic 
acid. 

For the relatively stronger acids that must be used for 
economical reasons, the advantage is with fluosilicic acid, 
both in amount of HF required and in cost of silicic acid as 
against boric acid. Thus a solution containing 5 gr. Pb and 
15 gr. BF 4 ' per 100 cc. has a resistance at 30° C. of about 
1.4 ohms, and a solution with 5 gr. Pb and 16.3 gr. SiF 6 " 
per 100 cc. (each containing 13.1 gr. F) has a resistance of 
about 1.3 ohm, per inch X inch 2 unit. 

Considering the higher cost of boric acid used as raw 



ELECTROLYTES FOR LEAD REFINING. 29 

material, these figures lead to the conclusion that fluosilicic 
acid is considerably the best. 

Fluosilicic acid is soluble in water and decomposable by 
alkalies into alkali fluoride and silica. Even as weak a base 
as litharge will effect a decomposition, which is the reason 
white lead and not litharge is used in making the lead salt. 
Heating causes a loss of acid by volatilization if the acid is 
strong. 

According to Baur,* Stolba noticed in 1863 that if fluo- 
silicic acid is boiled down the residue will dissolve silica, 
therefore SiF 4 must have escaped in boiling. The author 
(Baur) has found it to be the case that an acid containing 
13.3% H 2 SiF 6 gives a distillate also containing H 2 SiF 6 . 
Weaker acids give distillates with excess of HF, stronger acids 
with excess of SiF 4 . If then concentrated H 2 SiF 6 is dis- 
tilled partly, without silica being present, the residue should 
be caapble of dissolving silica; if weak acids 5-10% are 
evaporated, silica should deposit. This is found by experi- 
ment to be the case. The relative amounts of steam and 
hydrogen and silicon fluorides escaping are not given by the 
author. 

The specific gravity of fluosilicic acids is given in 
Table 11, taken from Comey's " Dictionary of Solubilities," 
originally given by Stolba. 

The preparation of lead fluosilicate solution from fluo- 
silicic acid can be successfully carried out in at least two 
ways. The most convenient method is to add lead carbon- 
ate or white lead, which dissolves with effervescence. 



* Berichte, Deutsch. Chem. Ges. 1903. 36 (16), 4209, abstracted Jour. 
»Soc. Chem., Vol. 27, page 17. 



30 



LEAD REFINING BY ELECTROLYSIS. 



TABLE 11. 



Per Cent H&iFe 


Specific Gravity. \ 


Per Cent H2SIF6 


Specific Gravity. 


2 


1.0161 


20 


1 . 1748 


4 


1 . 0324 


22 


1 . 1941 


6 


1 . 0491 


24 


1.2136 


8 


1 . 0661 


26 


1.2335 


10 


1.0834 


28 


1 . 2537 


12 


1.1011 


30 


1.2742 


14 


1.1190 


32 


1.2951 


16 


1 . 1373 


34 


1.3162 


18 


1 . 1559 







Iii his paper Mr. Semi * describes an experiment in 
which he added to 100 grams of 19.2% H 2 SiF 6 , 100 grams 
of lead as white lead, and got a precipitate containing 83.3% 
PbF 2 and 16.68% Si0 2 . This is of course the result when a 
great excess of lead is used, which is not. however, a matter 
of practical importance. Practically in making lead fluo- 
silicate solution, little or no precipitate is formed. 

Perhaps a cheaper method, though a less convenient one, 
is to elect rolyze the solution with lead anode and cathode, 
separated by a diaphragm. I made up about 10 cubic feet 
of solution experimentally, in fact this was the first method 
used. The solution was brighter and whiter and gave a bet- 
ter deposit on the start than that made in the other way. 
It also happened to contain an excess of HF. The solution 
was stored in carboys, and it finally dissolved the glass and 
ran out. Yet the excess of HF did not cause any precipi- 
tation of PbF 2 when the solution was used in refining. 

In using this method, the lead anodes dissolved evenly, 
the e.m.f. of the cell was about 1J volts, and no precipita- 



* Zeitschrift fur Elektrocliemie. April 14, 1905. 



ELECTROLYTES FOR LEAD REFINING. 



31 



tion was formed. A very little black spongy lead deposited 
on the cathodes, with much hydrogen. That the HF present 
did not precipitate lead fluoride, is due to the fact that HF 
is relatively a weaker acid than H 2 SiF 6 , and PbF 2 is not 
entirely an insoluble salt. The saving made by using lead 
as raw material instead of white lead, and apparatus to be 
used, are treated on pages 243 and 244. The apparatus used 
in making the solution is also shown in Fig. 1. 




Fig. 1, 

The crystallization of lead fluosilicate is a difficult mat- 
ter. The best results are got by placing a strong, nearly 
neutral solution over sulphuric acid under a bell jar and giving 
the solution several weeks to concentrate and crystallize, when 
beautiful crystals are obtained. The evaporation of the solu- 
tion even at 40-50° C. causes the precipitation of a fine crys- 
talline product of inexact composition, not entirely soluble 
in water. Crystals have also been got by dissolving lead 



32 LEAD REFINING BY ELECTROLYSIS. 

and lead peroxide in very strong fluosilicic acid and lead 
fluosilicate solutions, from two electrodes connected together 
through a resistance, 

Lead-fluosilicate crystallizes in very soluble, brilliant crys- 
tals, resembling those of lead-nitrate, and containing four 
molecules of water of crystallization, with the formula 
PbSiF 6 -4H 2 0. This salt dissolves at 15° C. in 28 per cent 
of its weight of water, making a syrupy solution of 2.38 sp. 
gr. Heated to 60° C, it melts in its water of crystallization. 
A neutral solution of lead-fluosilicate is partially decomposed 
on heating, with formation of a basic insoluble salt and free 
fluosilicic acid, which keeps the rest of the salt in solution. 

The electrolysis of fluosilicic acid and probably also of 
fluosilicates, is not entirely a simple electrolysis in which the 
ions H' and HSiF 6 ' take part. There is a tendency toward 
decomposition into Si0 2 and 6HF, the reverse of its forma- 
tion. Late experiments indicate that this takes place to some 
considerable extent, but for the most part the HF liberated 
at the cathode and the silica at the anode recombine under 
the influence of circulation and diffusion. An excess of HF 
in the solution would obviously tend to prevent the forma- 
tion of silica, and a solution containing excess of silica would 
deposit silica in the anode slime until a condition of equilibrium 
was arrived at, when no more silica would deposit. There 
is a certain loss of fluosilicic acid in actual practice which I 
regard is mostly due to mechanical loss by leaks, etc., because 
the silica in the slime is generally about 2% only, or about 



* Betts and Kern, Trans. Am. Electrochem. Soc, Vol. 6, page 6i 
Clarke, Am. Chem. Soc. 

t F. W. Clarke, Jr., Am. Chem. Soc, 28, 306, 190. 
J Private communication from the management. 



ELECTROLYTES FOR LEAD REFINING. 33 

one pound per ton of lead, corresponding to 2.3 pounds of 
H 2 SiF 6 decomposed. Solution has been thought to dissociate 
into which HF evaporates into the air, while the corresponding 
silica remains in the slime. The actual amount of HF present 
in the solution is usually slight, and its evaporation must be 
very small, on account of small vapor tension and high com- 
bining power with water. The fumes produced in a closed 
tank-room, refining perhaps 70 tons of lead daily, on the sup- 
position that the acid is lost in the air to the extent of 100 
to 300 or more lbs. fluorine in the form of SiF 4 and HF per 
day, would make the air unbearable, whereas the actual con- 
dition is that there is no noticeable acid fume in the air 
even in winter with the building closed. I cannot, therefore, 
believe that appreciable quantities of acid are lost by evap- 
oration from the tanks. 

There is always, of course, a considerable mechanical loss 
in the large bulk of slime, in the pores of the cathodes, and 
on the surface of both cathode and anode scrap, and from 
leaks in the tanks. New tanks absorb some solution and 
the salt PbSiF 6 probably crystallizes in the wood, which also 
causes a loss with new tanks. In view of these facts and also 
analyses of slime, the loss of acid by electrolytic decompo- 
sition not offset by the reaction between the Si0 2 and HF 
formed is probably extremely small. 

That silica deposits on anodes from solutions containing 
fluosilicic acid has been proved by electrolyzing solutions of 
ferric sulphate containing fluosilicic acids and analyzing the 
slimy coating on the anode in a similar experiment, and by 
the electrolysis of ferrous fluosilicate ;* in both cases with an 
insoluble carbon anode. 

* Private communication. Aug. E. Knorr. 



34 LEAD REFINING BY ELECTROLYSIS. 

In either case silica deposits on the anode, whereas if 
H 2 SiF 6 was not decomposable in solution no such thing would 
occur. In his article " Zur Kenntnis der Elektrolytischen 
Bleiraffination, " H. Senn also described an experiment in 
which he electrolyzed fluosilicic acid with platinum electrodes, 
wmen silica separated. 

" Experiment 40. — I used as electrolyte fluosilicic acid 
of specific gravity 1.267 (36.7 gr. H 2 SiF 6 per 100 cc). This 
contained a little hydrofluoric acid. Electrodes: platinum. 
Anode surface: 36.8 sq. cms. Cathode surface 41.6 sq. cms. 
Current: 0.45 amperes. Tension: 2.7 volts. Time 19 hrs. 

" At the close of the research the anode and the bottom 
of the glass were covered with a layer of gelatinous silica. 
The electrolyte had the peculiar smell of hydrofluoric acid. 
This had attacked the glass. Since I had no method of 
determining hydrofluoric acid in presence of fluosilicic acid, 
I had to be content with a qualitative proof. 

" The fluosilicic acid I filtered off and found in 250 cc. 
of electrolyte .1841 grams Si0 2 , corresponding to .4401 grams 
H 2 SiF 6 . This decomposition is indeed a result of the fact 
that SiF 6 was discharged on the anode." 

That the silica should deposit on the anode rather than 
on the cathode is a little surprising at first. If there is a 
dissociation of H 2 SiF 6 into HF and Si0 2 , as some have thought, 
HF is so much more a conductor than Si0 2 that it would 
apparently go to the anode to a greater extent than Si0 2 , 
and at the anode there would be an excess of HF, not Si0 2 . 

In the experiment of Senn's the proportion of silica de- 
posited corresponds to a decomposition of 0.48% of all the 
H 2 SiF 6 present, and as the glass was attacked, some or all 
of this must have come from the glass. In this experiment 



ELECTROLYTES FOR LEAD REFINING. 35 

there may have been a good deal more decomposition than 
this, as the continual circulation would bring the anode and 
cathode products together again, when the original equilib- 
rium from formation of H 2 SiF 6 would be again established. 

The maximum chemical and mechanical loss cannot be 
more than G lbs. of H 2 SiF 6 per ton of lead deposited, for 
analyses of a solution, thoroughly mixed both before and 
after a certain 150 tons of lead was deposited with a current 
density of 10-12 amperes per square foot, showed only this 
amount of loss. The solution contained 15% SiF 6 and about 
6% Pb. Any greater loss than this, observed w 7 hen working 
at this current density, must then be an avoidable mechanical 
loss, as indeed part of this 6 lbs. loss must have been. 

A sample of Trail slime from regular running, analyzed 
by me, contained 2.2% Si0 2 including silica in H 2 SiF 6 present. 
The slime of course is not completely washed over, and part 
or all of this silica then is due to the electrolyte not washed 
out. This shows a maximum loss in slime of about 2.1 lbs. 
H 2 SiF 6 per ton lead, occurring at the time that particular 
slime was made. Even part of the apparent loss is in some 
cases the result of deposition of excess of silica present 
in the solution used. 

Solutions of fluosilicic acid may contain excess of silica, 
and it is probable that H 2 SiOF 4 is formed to at least a slight 
extent, while after in use some time they probably contain 
excess of HF. In all cases the amount of the unstable com- 
pounds in solution will vary w r ith the concentration, tem- 
perature, etc. For that reason I think determinations of 
silica in anode residues from lead refined with new solutions 
are not reliable as indicating the extent to which H 2 SiF 6 is 
decomposed, nor what a lead fmosilicate solution will do after 



36 LEAD REFINING BY ELECTROLYSIS. 

it has practically reached its condition of equilibrium. IL 
Senn gives a few illuminating analyses in his paper which 
show the point I am making. 





TABLE 


12. 




Experiment Number. 








SiO-2 in Slim? 


28 








24.S7^ C 


30 








25.12% 


34 








1.82% 


35 








1.26% 


36 








•9% 



It will be noted that the large proportion of Si02 was 
found in slime from new solution. 

Starting with a solution which will deposit silica in the 
slime, while the ratio of Si to F in the solution gradually 
becomes less, ultimately a condition of equilibrium will be 
reached for any given constant conditions of temperature, 
current density, strength of solution, etc. 

Just at what point equilibrium will be reached, that is 
when the power of the solution containing free HForH 2 SiF 6 , 
capable of combining with silica to form H 2 SiF 4 or H 2 SiF 6 , 
is exactly balanced by the tendency of the current to deposit 
silica on the anodes is impossible to say, and on account of 
the constantly varying conditions it is not apt to be deter- 
mined. 

However, equilibrium will be reached long before the solu- 
tion can contain so much free HF that PbF 2 can precipitate 
or other unclesired reactions occur. With an ordinary solu- 
tion, say 7 grams Pb and 16 grams SiF 6 per 100 cc. as much 
as o% free HF may be present without precipitating PbF 2 . 
That is, the acidity clue to HF may be as high as that due 
to free H 2 SiF 6 without precipitating PbF 2 , of course on 
account of the fact that H 2 SiF 6 is relatively a much stronger 



ELECTROLYTES FOR LEAD REFINING. 37 

acid than HF, and is able to decompose a limited amount 
of insoluble PbF 2 . 

Evidently if a start is made with a particular solution, 
the slime may very easily contain a good deal of silica on the 
start, until equilibrium is reached, but this does not mean 
a loss of valuable fluorine, but of relatively valueless silica. 
On the other hand, by having a certain amount of free HF 
present, the slime can contain no precipitated silica. In 
practice very high silica in the slime, say 15%, is apt to be 
obtained on the start with a new solution. 

Evidently the surface of the cathodes, anode scrap, and 
metal particles of the slime that must be taken from the 
solution and wetted by it, is quite large. On account of the 
ease with which slime can be broken up and washed, the loss 
in the slime can be reduced to almost any extent. The sur- 
face of the anode scrap and cathodes can also be freed from 
acid to any extent by washing, but if there are any pockets 
of solution in the cathodes, or any chemical combination of 
lead with the electrolyte, as in copper deposited from the 
acetate solution under certain conditions,* such losses would 
be unavoidable. 

To investigate the losses practically resulting from elec- 
trolyte carried off by the cathodes, six pieces of cathodes from 
a large pile were obtained from the United States Metals 
Refining Company's Grasselli plant, and analyzed as follows: 
Samples of 100-150 grams were dissolved slowly while warm- 
ing only slightly in dilute nitric acid, of which only a rather 
small excess was used. A very little water-glass solution 
was added to the dilute acid on the start, to insure that the 

* "Ueber das Acetatkupfer. " Carl Benedicks. Metallurgie, 1907 (4) 5. 



38 



LEAD REFINING BY ELECTROLYSIS. 



fluorine present would be combined as SiF 6 . The water-glass 
dissolved entirely, but during the solution of the lead some 
silica separated. This was filtered off, the free nitric acid 
nearly neutralized with caustic potash (by alcohol) and a 
large excess of potassium nitrate and acetate added. The 




Plate 1. — Samples of Lead Cathodes. 

precipitated K 2 SiF 6 was filtered off and titrated, giving the 
following results. The photographs show the lead pieces from 
which the samples were cut, one photograph showing one 
side and the other photograph the other side. 



TABLE 13. 



Number in 
Photograph. 

1 015% SiF 6 

2 011% " 

3 005% " 

4 009% " 

5 002% " 

6 003% " 



0.30 lb. SiF 6 per ton lead. 

0.22 

0.10 

0.18 

0.04 

0.06 



ELECTROLYTES FOR LEAD REFINING. 



39 



We have evidently a very small loss of acid in the cathodes, 
and I have been informed that results similar to mine have 
been obtained at the Trail refinery.* 

The method of washing cathodes in use when the 4 above 
pieces were made was to wash them with water first, and use 




Plate 2 — Samples of Lead Cathodes. 

the wash-water over and over until nearly of the same strength 
as the main electrolyte, when the wash-water is added to 
the electrolytic tanks. The average loss from acid solution 
remaining on the cathodes after draining is evidently about 
one-half the loss involved if the electrolyte was merely allowed 
to drain off. To determine just what this loss might be, 
samples 2, 4, 5, and 6 in the photographs, were cleaned of 
a surface coating of white lead, and weighed after wetting 
and draining for a minute or two, and again after becoming 
dry, with the result given in Table 14. 

* Letter from Mr. W. H. Aldridsre. 



40 LEAD REFINING BY ELECTROLYSIS. 

TABLE 14. 



Number in Weight Cath- c-i .• „ Acid Loss 

Photo- ode per *°^Z. per Ton 

graph. Square Foot. «- athode - ; Lead. 



28.8 lbs. 0.50% 



1.66 lbs. 
SiF fi 



Average 
Loss in 
Practice. 

0.83 lbs. 
SiF fi 



lbs. , 0.39% 



1.33 lbs. 0.67 lbs. 
SiF, SiF fi 



16 



lbs. 0.36% 1.20 lbs. 
SiF. 



0.60 lbs. 

SiF, 



11 



lbs. 



0.22% 



0.76 lbs. 
SiF., 



0.38 lbs. 

SiF fi 



Remarks. 



Cathode average 
weight and 
roughness. 



Cathode average 
weight and 



roughness. 



Unusual cathode. 



Unusual cat ho.! 
not well wetted. 



The above cathodes were deposited from a solution con- 
taining 8% Pb, and are solider than those obtained with a 
solution containing 6% Pb, as used at Trail, which may 
account somewhat for the higher acid loss at Trail. 

The maximum loss, with fairly solid cathodes of average 
thickness, of 28 to 30 lbs. per square foot, is not greater than 
1 lb. SiF 6 per ton lead. The loss on anode scrap cannot be 
over 30 to 40% of this amount, on account of smaller surface 
(1 anode makes 2 cathodes usually) and smoother surface both. 

The loss in slime may be reduced by a moderate amount 
of washing to 2 lbs. SiF 6 per ton lead or lower. The loss out- 
side of leaks, which can of course vary extremely, and evap- 
oration from the tanks which cannot be otherwise than neg- 
ligible, when the air in the tank-room has no acid smell, as 
is usually the case, cannot then be over 3.5 lbs. SiF 6 per ton. 
If more than this, something is wrong with the plant, which 
might be due to a bad leak or not sufficient washing of the 
slime, or deposition of soft rotten lead on the cathodes or 
other less evident causes. 



ELECTROLYTES FOR LEAD REFINING. 



41 



The acid loss at Trail in 1902 and early in 1903 was as 
given in Table 15. 

TABLE 15. 



Aug. 3— Sept. 16, 


245 tons deposited 


13.8 lbs. SiF 6 per ton deposited 


So ju. 16- -Oct. 6, 


120 


ij ij ( ( it i ( c e ii 


Jan. 22— Feb. 13, 


135-145 " 


6.3 " " " " 



The solution was, however, weaker than is used at present, 
as follows: 

TABLE 16 



Aug. 3, 


7.86% Pb 


10.58% SiF c 


Sept. 16, 


6.19% " 


7 . 94% ' ' 


Oct. 6, 


6.07% " 


6 . 93% ' ' 


Jan. 17, 


6.40% " 


8.56% " 



On the other hand during at least the first two of the 
above periods, no evaporation of wash-water was practised, 
and only enough was used in a crude way to make up for 
the solution taken out. There were leaks, too, and no suit- 
able apparatus for catching a good share of them. 

The average amperes and volts at the time may be seen 
from Table 17. 



Am 



Sept. 



Oct. 



Nov 



TABLE 17. 

Current 
Efficiency. 

1-15 3393 amps. .293 volts per tank 59 % 12.8 amps, per sq. ft. 



Average. 



15-31 3196 
1-15 3406 

15-30 3148 
1-15 2724 

15-31 2593 
1-15 2247 

15-30 1891 



.328 

.39 

.42 

.44 

.435 

.435 

.42 



90 % 12.1 
72i% 12.9 
74 % 11.9 
89i% 10.6 

92 % 9.8 
81£% 8.5 

93 % 7.2 



The apparent heavy loss for the first period was probably 
due to absorption by the new tanks, leaks, and the unsettled 
condition of everything, but principally the solution was 
not well mixed, for the sample indicates considerably more 
acid than was actually purchased by the plant. The loss 



42 LEAD REFINING BY ELECTROLYSIS. 

for the third period, with a weaker electrolyte, however, 
shows better work than is reported at present. Before Janu- 
ary 17th some new acid had been added. It would appear 
that the high acid makes much higher acid loss, but that con- 
clusion is not safe, as other conditions were changed very 
much during the first few months. 

The actual loss of acid experienced at the Trail refinery 
up to the present, I have been informed by the management, 
is about 10 lbs. SiF 6 per ton lead. Probably this has been 
improved since the figure was determined some time ago. 

Tables 18 and 19 give the electrical resistance of acid lead 
nuosilicate solutions. Table 18 is from determinations by 
Dr. E. F. Kern in my laboratory. The other table gives older 
determinations made by myself, and includes many solutions 
of no practical importance, but at the time the table was 
made it was not known which solutions would be most de- 
sirable. The temperature coefficients are obtainable from 
Table 19. The conductivity of the solutions are also plotted 
as Figs. 2, 3, and 4. 

The amount of gelatine required under good working con- 
ditions is not great, and may be taken at from J lb. to f lb. 
per ton of lead deposited. Gelatine in She form of glue is 
always used, as it is cheaper. I believe the better grades of 
glue the most suitable, for some of the cheapest glue makes 
a disagreeable smell in the tank-room. In practical work, 
when the glue in the solution is about used up, and it is 
necessary to use more, there will be noticed on the cathodes 
a tendency toward the formation of points on the lumps, 
which are readily noticeable with a little practice. The glue 
is added in the form of a hot, strong solution, and may be 
best put in the circulation-tank a little at a time. 



ELECTROLYTIC FOR LEAD REFINING. 



43 



The appearance of the pure lead fluosilicate solution is 
that of a colorless liquid. After it has been in use for some 
time, it has sometimes acquired a greenish tinge from traces 
of iron and perhaps nickel in the lead anodes. If the solu- 
tion is allowed to stand in contact with air away from the 
reducing action of the electrodes, it acquires a brownish yel- 
low color, which at first was thought to be a ferric salt, but 
now I believe it is due to a coloring-matter introduced with 





il2Pb 
\ 






4P 


b 
\ 8Pb 


N. 



























12 16 

Grams Si F 6 per 100 c. c. 

Fig. 2. 



20 



23 



the glue. On again using the solution for refining it becomes 
colorless again, due probably to the reduction of the coloring- 
matter to the reduced or "leuco" condition, winch is a 
characteristic of most organic coloring-matters. 

The metallic elements that enter into consideration as 
possible constituents of the electrolyte are the elements usu- 
ally present in lead bullion, those that may be in the fluo- 
silicic acid as impurities at the start and the iron binding 



44 



LEAD REFINING BY ELECTROLYSIS. 





11 12 28 24 30.5 

Grams Si F 6 per 100 c. c. 10 Grams Pb per 100 c. c. 

Fig. 4. 



ELECTROLYTES FOR LEAD REFINING. 
TABLE 18. 



45 



KFa lira ins 
per 100 cc. 


Lead Grams 
per 100 CO. 


Temperature. 


Resistance. 


Resistance 

at 20° C. 

Calculated. 


23.3 
23.0 

23 . 
23.0 
23.0 


23.6 

16 

12 

8 

3.4 


19. 5° C. 
17. °C. 

19. °C. 

20. °C. 
20. °C. 


2.34 
1.76 
1.17 

1.09 
.87* 


2.31 
1.60 
1.12 
1.09 
1.06 


16.0 
16.0 
16.0 
16.0 


16.0 

12.0 

8.0 

4.0 


16. °C. 
19. 5° C. 
20. °C. 

20. °C. 


2.68 
2.07 
1.49 
1.31 


2.56 
2.05 
1.49 
1.31 


12.0 
12.0 
12.0 


12.0 
8.0 
4.0 


15. 5° C. 
20. °C. 
20. °C. 


3.59 
2.32 
1.63 


3.24 
2.32 
1.63 


8.0 
8.0 


8.0 
4.0 


15. °C. 
19. °C. 


4.69 
2.79 


4.19 
2.73 


4.0 
63. Of 


4.0 

85.0 


13. °C. 
20. °C. 


9.22 
4.84 


8 
4.84 



* Incorrect. Correct figures is about 1.06. 

f Saturated solution PbSiF fi , specific gravily, 2.32. 



TABLE 19. 



SiF6 Grams 
per 100 cc. 


Lead Grams 
per 100 cc. 


o e c. 


10° C. 


20° C. 


30° C 


30.5 


27.8 


2.95 


2.18 


2.10 


1.84 


30.5 


25 


2.66 


2.15 


1.84 


1.60 


30.5 


20 


2.07 


1.72 


1.57 


1.21 


30.5 


15 


1.74 


1.45 


1.23 


1.07 


30.5 


10 


1.48 


1.21 


1.04 


.75 


27.1 


25 


3.22 


2.49 


2.13 


1.86 


24 


20 


2.73 


2.24 


1.72 


1.52 


24 


15 


2.01 


1.67 


1.33 


1.14 


24 


10 


1.62 


1,33 


1.14 


.99 


24 


5 


1.31 


1.45 


1.14 


.87 


21.9 


20 


3.39 


2.68 


2.32 


1.99 


18 


15 


3.50 


2.99 


2.34 


2.09 


18 


10 


2.13 


1.77 


1.50 


1.31 


16.4 


15 


3.80 


3.25 


2.54 


2.28 


12 


10 


4.62 


3.74 


3.35 


2.69 


11 


10 


4.84 


4.13 


3.51 


2.81 



46 LEAD REFINING BY ELECTROLYSIS. 

of the tanks. The elements being considered then are iron, 
zinc, sulphur, copper, nickel, tin, antimony, arsenic, silver, 
bismuth, cadmium, gold, selenium, tellurium, and other ele- 
ments in smaller quantities. Of these antimony, arsenic, 
silver, gold, copper, bismuth, selenium, tellurium are easily 
precipitable b} r lead, and consequently if they get into the 
solution they will be thrown out by the lead electrodes, 
mostly by the cathodes I believe, for the anodes are usually 
covered with slime, which would prevent their reducing, for 
instance, much antimony, although the antimony of the 
slime ' would quickly enough throw out such an easily pre- 
cipitable metal as silver. 

Zinc, iron, and nickel, if they find their way into solu- 
tion remain there, as they are not precipitable by the lead 
electrodes, nor can they be in any way thrown out on the 
cathode by the electric current so long as there is a fair amount 
of lead in solution, which there always is. Analyses of lead 
bullion show the presence of iron and zinc in small quanti- 
ties, say .02%. Whether the iron really does dissolve, I 
doubt, because the slime usually contains from one-half to 
two per cent of iron, accounting for at least a considerable 
part of it. The slime also contains sulphur as sulphides. As 
iron is not liberated in the lead smelting-furnace, but only 
iron sulphide, the lead then probably takes up small amounts 
of matte, and perhaps contains lead sulphide too. 

It is difficult to see how any of the sulphur of the lead 
bullion could get into the solution, except possibly as H 2 S. 
At any rate no sulphur has yet been either observed in the 
solution nor found in the lead by analysis. 

The other element, tin, occupies practically the same posi- 
tion in the scale of electromotive forces of solution that lead 



ELECTROLYTES FOR LEAD REFINING. 47 

does. That is to say, it takes about the same electromotive 
force to deposit tin (from the acid solution) that it does lead, 
or to dissolve it from the anode. Consequently a mixture 
of tin and lead can behave as practically one metal. 

Dr. Hans Mennicke * made some researches on the refin- 
ing of tin lead alloys, with a solution of tin fluosilicate. The 
solution was more difficult to prepare than the lead solution. 
No gelatine was added, but he produced good tin deposits 
so long as only a little or no lead was in the solution. With 
anodes of solder his deposits soon got spongy. With gelatine 
added, the deposits would probably have been solid. Both 
tin and lead dissolved from the anodes. Dr. Mennicke's re- 
sults were not successful, but not conclusive either. 

With about .02% tin in the anodes at Trail, some tin went 
over to the cathodes. After its presence, which was not sus- 
pected at first, was proved, it was removed by poling the 
lead before casting when the dross on the lead took up the 
tin. In such cases the dross could be smelted to lead con- 
taining tin, and the tin recovered by the usual softening pro- 
cess practiced as a preliminary in refining lead by the Parkes 
process. The percentage of tin found in the dross can be 
calculated on the basis that 4 or 5% of dross is produced. 
The analyses show: 

Bullion. Lead before Poling. 
Average 0289% Sn Average 0063% Sn 

The analyses indicate that either part of the tin remains 
in the slime or part has already got into the dross before 
poling. The latter must have certainly taken place and prob- 
ably the former to some extent too. In refining some bullion 

* Elektrochemische Zeitschrift, Vol. XII, 112, 134, 161, 180. 



4S LEAD REFINING BY ELECTROLYSIS. 

with about 4% tin, very considerable quantities of tin remained 
in the slime. The tin in such slime dissolved out with evo- 
lution of hydrogen on treating with HC1. 

A bar of fine solder (60 tin, 40 lead) I once made anode 
in a lead fluosilicate solution containing gelatine. The anode 
dissolved regularly and the cathode deposit was excellent. 
The experiment was not concluded. 

Tin is very rarely found in lead bullion, but in case it is, 
part or all can be recovered from the dross produced in melt- 
ing the cathodes, and the remainder in the slime can be re- 
covered in several ways, if the quantity is large enough to 
make it pay. 

The nature and influence of the anode slime. — Almost 
invariably in practical refining the anode slime remains 
attached to the anode, and very little change in appearance 
is noted, even when the lead is nearly all gone. This is espe- 
cially the case when the lead contains a considerable amount 
of antimony, say 109c > when there is hardly any change in 
the color even. The slime is of varying degrees of hardness. 
The proportionate spaces occupied by the metal of the slime 
and the liquid with which it is saturated can be calculated 
with some accuracy. The data are: 1 cubic centimeter of 
lead weighs 11.36 grams, of antimony 6.7 grams, of copper 
8.9 grams, and of silver 10.5 grams. Allowing 10% of lead 
in the slime, the actual space occupied by the slime from 
11.36 grams of alloy (specific gravity practically the same as 
that of lead) is ordinarily about .035 cc, leaving .965 cc. for 
solution, or in percentages the solution occupies approxi- 
mately 96-97^c and the slime 3-4%. The metal of the slime 
is prevented from carrying any of the current passing through, 
at any rate after it has reached its final composition,, by 



ELECTROLYTIC FOR LEAD REFINING. 



49 



polarization. To illustrate, suppose there is a piece of some 
metal lying below lead in the electromotive force series for 
fluosilicic acids, as copper, in the electrolyte between a lead 
anode and cathode, Fig. 5. 

Instead of passing through the copper, by reason of its 
greater conductivity, the current passes around the copper 
on account of its polarization. Current can only pass to the 




Fig. 5. 



copper by depositing lead on it. On the other side where 
the current leaves, something else would have to dissolve, 
and this must be copper. To deposit lead on one side and 
dissolve copper from the other would require an electromo- 
tive force of about 0.5 volt. This is not the actual truth of 
the matter, because very small currents can pass to and from 
electrolytes and electrodes without any evidence of electrol- 
ysis. With more than an extremely small current the cur- 



50 



LEAD REFINING BY ELECTROLYSIS. 



rent passes around the copper as if it were a piece of 
glass. 

If now the current is increased so that the fall of poten- 
tial in a distance about equal to the diameter of the copper 
piece approximates 0.5 volt, current will begin to go through 
the copper to a considerable extent, while copper goes into 
solution from one side. 

The slime in practice consists of a number of different 
metals of course, antimony, arsenic, bismuth, copper, silver, 
and gold, with some combined lead. For the acid fluosilicate 
solution, the electromotive force series has been determined 
by dipping a piece of lead and of one of the other metals into 
the lead electrolyte, by Dr. Kern. 



Pb 


TABLE 


20. 


.0 
+ 43 


volts 


Zn 








Al 






_u 


05 
03 
08 
37 
42 
43 
52 
60 
63 
68 


< < 


Sn 






I ( 


Fe. . 






< ( 


Sb 






1 I 


Bi 






a 


Cu. ... 






t ( 


As 






( c 


Ag 






t t 


Pt 






,< 


C 






a 



Such a method, however, cannot be accepted as giving 
the correct figures. A better way to determine these values 
is to prepare two fluosilicate solutions, one of the lead and 
one of the other metal, separate the solutions with a porous 
partition, dip the two metals in their respective solutions, 
and read the e.m.f. with a suitable instrument. 

Somewhat different and more correct figures have been 



ELECTROLYTES VOR LEAD REFINING. 



51 



obtained by placing load fluosilicatc solution both inside and 
outside of a porous cell, contained in a beaker. Such a cell 
as that shown in the sketch (Fig. G) is useful and handy for 
Buch purposes. 

It consists of a small piece of wood, which is heated in 
paraffine, and after cooling, pieces of paper, asbestos or cloth 
are cemented on the wood with warm paraffine. In this par- 
ticular experiment I used ordinary cardboard as diaphragm. 




Fig. 6 

A piece of the metal being tested is hung in the inside 
solution with a platinum wire and a piece of lead dips in the 
outside solution. A small current is passed for perhaps five 
minutes with the metal under investigation as anode. On 
shutting off the current the e.m.f. is read with a millivoltmeter. 
It is difficult to get a constant reading with arsenic, so I took 
the highest observed. I think the arsenious fluosilicate formed 
at first goes quickly over to arsenious acid, which gives a 



52 LEAD REFINING BY ELECTROLYSIS. 

lower e.m.f. against lead. As the mechanism of the reaction 
seems to be first the formation of an arsenious salt, the high- 
est e.m.f. is the one with which we are concerned when in- 
vestigating the possibility of arsenic dissolving with the lead. 
The series is given in Table 21. 

TABLE 21. 

Lead volts. 

Arsenic 40 " 

Antimony 43 " 

Bismuth 47 " 

Copper 51 " 

Silver 97 " 

These figures represent an approximately correct series 
for most other oxygen-acid electrolytes. According to it 
each metal in the series cannot be precipitated by those pre- 
ceding, while if the difference of e.m.f. is at all considerable 
(say 0.1 volt), each metal will precipitate those following. 
When the difference is, however, so small as between anti- 
mony and bismuth, no precipitation is usually observed to 
take place on dipping the higher metal into a solution of the 
next lower one. However, by the electrolysis of a mixed 
solution, for example, copper and bismuth methyl-sulphate, 
copper can be deposited out with little or no bismuth. 

As we have seen already, with ordinary lead, such as con- 
tains say 3% of impurity, the space actually occupied by the 
metal of the adhering slime is only about 3% of the total. 
Under the conditions of electromotive force existing, this 
metal is practically non-conducting, on account of polariza- 
tion. Its bulk is, however, too small to directly affect the 
conductivity of the solution with which it is saturated. It 
does have some small effect, however, by obstructing free 
circulation of the electrolyte in the neighborhood of the anode, 



ELECTROLYTES FOR LEAD REFINING. 53 

but the total resistance 4 thus introduced is extremely small. 
In actual refining, there is rather a tendency for the electro- 
motive force to fall off than to increase, as the electrolytic 
action on a set of anodes and cathodes goes on. 

The usual thickness of an anode is about one inch, and 
the maximum thickness of slime then is about one-half inch. 
Any current so large or conductivity so low that the drop 
of potential in traversing one-half inch amounts to .4 volt, 
or .8 volt per inch, would be as capable of dissolving arsenic, 
antimony, bismuth, and copper from the surface layer, as 
of dissolving lead from the solid electrode beneath. With 
an electrolyte having a resistance for the cubic inch unit of 
1.4 ohms, the maximum current strength per square foot per- 
missible would be .8 X^X 144 = about 82 amperes. As this 

is far beyond any current that it is practicable to use, there 
is' never any clanger of contamination of the solution or the 
refined lead from the direct attack of the anodes by the cur- 
rent, except in the case of tin. This statement applies only 
to lead alloys with a largely preponderating proportion of 
lead. 

Regarding the nature and constitution of the slime, using 
as anode alloys of lead and various other metals, it makes 
some difference in the amount of lead retained by the slime 
what the other metal is, and how it is combined with the 
lead, and the speed of working appears to have some influ- 
ence, high current density leaving more lead in the slime than 
low-current density. There appears to be no appreciable 
combination between copper and lead. Alloys of copper and 
lead have given a slime practically free from lead. Alloys 
with 40% copper and 60% lead can be treated easily. Experi- 



54 LEAD REFINING BY ELECTROLYSIS. 

merits were made by Dr. Kern, which are described in 
Chapter IX. 

Silver also holds back very little lead. For example silver- 
lead alloys of composition given in Table 22. 





TABLE 


22. 




Pb 


88% 




82.37% 


Ag 


0.75% 




14.60% 


Cu 


1.53% 




2.22% 


Sb 


•5% 




•77% 


Bi 


1.11% 




.19% 



gave pure lead readily and slime containing only 1.5% Pb 
and 2.1% Pb, respectively. Very likely what little lead was 
left was in combination with the antimony and bismuth. 

That lead does combine with some metals to form com- 
binations not decomposed on electrolyzing the alloy as anode, 
is very forcibly brought out by Mr. Senn's experiment on the 
refining of lead-platinum alloy.* The alloy contained 10.10% 
platinum, and the slime contained 70.45% Pb and .16% Si0 2 . 
In another experiment the alloy contained 10% platinum 
and left a slime of fine leafy crystals, containing Si0 2 1.08%, 
Pb 65.30%, and Pt 32.93%, corresponding to the formula 
PtPb 2 . Even when the voltage was raised so high as 3 volts 
with oxygen evolution at the anode, the compound did not 
decompose to any great extent. Lead peroxide was formed, 
but that probably was deposited from the solution. 

Bismuth probably retains about one-sixth of its weight 
of lead, and antimony J to i of its weight. The amounts 
can, however, be very different according to the current den- 
sities used. 

* Senn, Zeitschrift fur Elektrochemie, April 14, 1905. 



ELECTROLYTES FOR LEAD REFINING. 



55 



The truth seems to be that the compounds of lead and 
other metals are decomposable, but only slowly. The slower 
the treatment the less lead in the slime. The maximum 
electromotive force available in practical refining for the 
decomposition of these compounds cannot well be over 0.05 
volts on the average and 0.15 volts as a maximum, and 
probably some of the compounds with a small proportion 
of lead are able to resist this. 

For the electrochemical decomposition of antimonides of 
lead, etc., the following experiment which I made is interest- 
ing: An anode of hard lead with about 18.8% antimony 
was used. Anode area 48 square inches. 



TABLE 23. 



Time. 


Hours Run. 


Current, 
Amperes . 


Volts. 


Current Density, 
Amperes per 
Square Foot. 


Back E.M.F. 


8.00 




13 




39 




8.30 




10 


.61 


30 




10.30 


2.5 


10.5 


.73 


31.5 




11.30 


3.5 


10.5 


.73 


31.5 




1.30 


5.5 


10.5 


.76 


31.5 




3.10 


5.5 


10 


.77 


30 




4.00 


6.3 


9 


.99 


27 


.232 


4.45 


7.0 


8 


.92 


24 




5.00 


7.5 


5.5 


.56 


16.5 


.232 


9.30 


24.0 


1.5 


.53 


4.5 


.304 



The back e.m.f. was the voltage read with the voltmeter 
on interrupting the current, and is a measure of the chemical 
affinity of the antimony of the slime for more lead than it 
is already combined with. Analysis of the residue gave 7.83% 
lead. 

The same anode was further electrolyzed with from 1 to 
2 amperes, when the back e.m.f. finally rose to .328 volts, 



56 LEAD REFINING BY ELECTROLYSIS. 

and the residue became so fragile that it broke. The lead 
antimony compound was then nearly entirely decomposed. 
As the difference of e.m.f. of solution of lead and antimony 
is about .43 volts, the e.m.f. still falls short of that necessary 
to dissolve antimony. 

The heat of combination of lead with excess of antimony 
is then about 17,000 cal. 

Since the maximum voltage available in refining ordinary 
lead bullion for decomposition of antimonides is only about 
.10 to .15 volts, the antimonide of lead cannot be nearly 
completely decomposed in usual practice. 

Pure lead was deposited by Dr. Kern in my laboratory 
from alloys of composition given in Table 24. 



Pb 65.37% 

Bi 7.32% 

Sb 19.51% 

As 5.85% 

Ag 1.95% 

Cu 

It was necessary to work with a low-current density of 
about four amperes per square foot with the first two alloys, 
and use thin anodes, -that would last say, six days, or what 
amounts to the same thing, clean them every six days. The 
slime from the first two alloys contained 5.30% Pb beside the 
bismuth, silver, copper, antimony, and arsenic. With the 
second two alloys, a continuous run was made without clean- 
ing anodes of ten days, back e.m.f. at end — .07 volts. 

For an idea of the amounts of lead held back by antimony, 
copper, and bismuth, the table from Mr. Senn's paper is in- 
structive. 



TABLE 24. 






65.56% 


82.79% 


88.52% 


6.94% 


3.42% 


2.28% 


18.24% 


9.12% 


6.08% 


5.47% 


2.73% 


1.82% 


1.94% 


.97% 


.68% 


1.94% 


■97% 


•68% 



ELECTROLYTES FOR LEAD REFINING. 
TABLE 25. 



57 



No. 


Anode Contain 


s Amperes. 


Current 
Amperes. 
Sq. Dm. 


Amperes, 
Sq. Ft. 


Dura- 
tion, 
Hours. 


Lead 

De- 
posited, 
Grams. 


Quantity 

Slime, 
W. 


Deposited 

Lead 
Contains 


27 


.92% Cu 


.5 


.59 


5.7 


29 


55.9 


1.64 


No Cu 


28 


.92% Cu 


.9 


1.07 


10.2 


24 


83.3 


.87 


NoCu 


29 


1.006% Cu 


1.3 


1.55 


14.9 


18 


90.2 


3.11 


NoCu 


30 


1.006% Cu 


2 


2.30 


20.2 


9 


69.4 


1.10 


NoCu 


31 


12% Bi 


.5 


.59 


5.7 


24 


48.3 




NoBi 


32 


12% Bi 


.9 


1.07 


10.2 


7.25 


25 


12.72 


NoBi 


33 


12% Bi 


1.3 


1.55 


14.9 


11 


55.1 


12.60 


NoBi 


34 


26 . 67% Bi 


.9 


1.07 


10.2 


17 


59. 


13.25 


No Bi 


35 


26 . 67% Bi 


1.3 


1.55 


14.9 


16.5 


82.6 


36.38 


.94% Bi 


36 


10.03% Sb 


.5 


.59 


5.7 


30 


57.9 


8.58 


NoSb 


37 


10.03% Sb 


.9 


1.07 


10.2 


8.5 


29.5 


5.19 


NoSb 


38 


10.03% Sb 


1.3 


1.55 


14.9 


2 


110.3 


18.52 


.13% Sb 


39 


9.81% Sb 


1.3 


1.55 


14.9 


18 


90.3 


10.43 


.05% Sb 


45 


10.01% Pt 


.28 


.59 


5.7 


16 










Analysis of Slime. 




No. 




Solution. 




Pb% 


Cu% 


Bi % 


Sb % 


Si0 2 % 


F% 




27 




23.41 








9% Pb— 11% free 


28 




36 


31 






28.47 




H 2 SiF 6 falling to 


29 




19. 


47 










4.83% Pb— 8.56% 


30 


io.03 


57. 


96 










25.12 




free acid. 


31 


70.4 


9 








32 








42.46 










33 








35.44 










34 


i2.4 
34.83 






83.97 
60.15 






1.82 
1.26 


9 




35 






36 


67.7 


1 


9% Pb— 11% free acid 


37 










47.52 


1.04 




falling to 2.48% Pb— 


38 










53.47 






11.5% free acid. 


39 










45.00 








45 





























The actual amount of lead thus retained with anodes of 
ordinary grades of lead bullion is quite small. If 60 pounds 
of slime are produced per ton of lead, and it contains 12% 
lead, which is a fair average, the amount of lead in the slime 
is 7.2 lbs. = 0.37% of the total. 
The electrolysis of the fluosilicate solution with an insol- 



58 LEAD REFINING BY ELECTROLYSIS. 

uble anode is interesting. In this case lead deposits on the 
cathode and lead peroxide on the anode leaving fluosilicic acid. 

2PbSiF 6 + 2H 2 = Pb0 2 + Pb + 2H 2 SiF 6 . 

This reaction may be useful, for instance if it becomes 
necessary to reduce the percentage of lead in the solution 
for any reason. Storage batteries have been constructed 
to work on this principle, but there are mechanical difficul- 
ties which have yet to be overcome. If a purification of 
refining solution became necessary the lead could be removed 
in this way and the solution distilled or purified in other 
ways. 

From the analogy to electrolytic copper refining the ques- 
tion of purifying the electrolyte was early given a good deal 
of consideration, and a number of purifying schemes proposed. 
We now know that the question of purification of solutions 
will never come up in refining ordinary grades of bullion. The 
only metals of the anodes that can accumulate in the elec- 
trolyte are iron, zinc, nickel, and cobalt. Most of the iron, 
which is very small in amount, remains in the slime and the 
others are only present in traces. The total amount of these 
metals dissolved is probably not over .01%. The loss of elec- 
trolyte, which will be .3 cubic foot or more per ton lead 
refined, permits a maximum of iron, etc., assuming .01% to 
dissolve of at most 1.07% of these metals in the solution, 
which is too small to be serious. 

As the preparation of pure lead may be of interest, the 
following quotation from a paper by Dr. Kern and myself 
on the "Lead Voltameter,"* is given: 

* Trans. Am. Electrochemical Society, Vol. VI, page 67. 



ELECTROLYTES FOR LEAD REFINING. 59 

" The solution was diluted so as to contain 17 grams of 
PbSiF 6 and 7 grams of free H 2 SiF 6 in 100 cc. solution. After 
adding one gram of gelatine (dissolved in hot water) to 2000 
cubic centimeters of solution, the electrolyte was rendered abso- 
lutely pure, in respect to metals which can deposit with lead, 
by electrolyzing for several days, using electrodes of refined 
lead. The anodes were wrapped with two thicknesses of 
clean linen, so as to prevent the impurities from dropping off 
and floating in the electrolyte. The small amount of solu- 
ble impurities in the electrolyte was clue principally to the 
impurities in the white lead used for making the solution. 
The electrolysis was continued for four days at temperatures 
between 17° C. and 57° C, using a current density at the 
electrodes of 10 to 12 amperes per square foot. A small 
amount of 'anode sludge' remained behind, and in order 
to prevent it from being oxidized by the atmosphere and 
subsequently going into solution, melted vaseline was poured 
on the surface of the electrolyte. The deposit which formed 
on the cathode was smooth, dense, and non-crystalline. 

" After purifying the electrolyte, about 800 grams of abso- 
lutely pure lead was made by electrolysis, using ordinary 
refined lead, wrapped with clean linen, for the anodes. The 
refined lead which was deposited on the cathode was fur- 
ther refined by reversing the current density and re-deposit- 
ing it on new cathodes. The solution was protected from 
the atmosphere by a covering of melted vaseline. The puri- 
fied lead was melted, cast into a thin plate, and then rolled 
into sheets about A inch to T V inch thick. The sheets were 
cut into strips of suitable size and used as anodes for the 
lead voltameter. No residue was left on dissolving the puri- 
fied anodes by electrolysis." 



CHAPTER II. 

CHEMISTRY OF SLIME TREATMENT. 

Lead slime contains originally metallic lead, copper, gold, 
silver, bismuth, antimony, arsenic, sulphur, and occasionally 
probably other elements, tin, selenium, and tellurium. For 
analyses see Table 26. The object of its refining is to recover 
especially the gold and silver, but the bismuth, antimony, 
copper, and lead are also valuable and should be saved. 

There are several different methods of treatment, based 
on different chemical or physical properties of the various 
metals. 

These methods are distillation, amalgamation, fusion to 
alloy, followed by chemical or electrochemical treatment of 
the alloy, fusion to bullion and slag, fusion to matte and slag, 
electrolytic refining of the slime direct, dry treatment with 
chlorine and separation of the chlorides by distillation, and 
various wet chemical and electrochemical methods of treat- 
ment. 

Distillation. — The temperatures necessary to distil the 
metals could easily be obtained in an electric furnace, and 
nothing would be simpler than to separate the metals in this 
way, apparently. The energy requirement would not be 
large. The trouble is that the boiling-points of the three 
principal metals, lead, antimony, and silver, lie too close 
together. That antimony boiled as high as 1600°, as deter- 

60 



CHEMISTRY OF SLIME TREATMENT 



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62 LEAD REFINING BY ELECTROLYSIS. 

mined by V. Meyer, did seem surprising, as antimony is re- 
garded in smelting as a volatile element. Its volatility in 
smelting is due to the low boiling-point of its oxide, how- 
ever. I was unable to volatilize any antimony on heating 
it in a carbon crucible in an anthracite fire, and the tem- 
perature was certainly high enough to melt steel. 

Messrs. Moissan and Watanabe * report the results of experi- 
ments on the distillation of alloys of nearly equal parts cop- 
per and silver; and of tin 64% and silver 35%, and of lead 
53% and silver 46%. The alloys were heated for various 
periods in an electric furnace and the residual metals weighed 
and analyzed. I have plotted the results as Figs. 7, 8, and 9. 
The last two figures are based on only three determinations 
each, and the curves are in consequence not necessarily en- 
tirely correct in form. They show no complete separation 
of metals by distillation. 

Amalgamation. — Lead, copper, gold, silver, and bismuth 
amalgamate easily, and arsenic and antimony do not, so that 
a separation might be made. Fresh unoxidized slime has 
been ground with mercury and some fluosilicate electrolyte 
in a mortar, and the mercury takes up a portion of the sil- 
ver, gold, copper, and lead, but the separation is very far 
from complete, probably because the metallic arsenides, anti- 
monides, and other compounds present in the slime are too 
stable to be decomposable by mercury. Even after the sep- 
aration was made by amalgamation, considerable still remains 
to be done before the metals are finally recovered. The bul- 
lion could be retorted as it is usually done in amalgamation 
silver mills, or an electrolytic method for extracting silver, 

* Comptes Rendus 1, CXLIV, 1907. Number 11, page 16. 



CHEMISTRY OF SLIME TREATMENT. 



63 



copper, lead, and bismuth from the amalgam might be 
devised. 








75.7$ 



Fig. 8. 




39.8jf 4Q.3ji 



Fig. 9. 



85.9* 



Fusion to alloys. — Slime oxidizes rather readily when 
dry, and some varieties will inflame spontaneously on drying, 



64 LEAD REFINING BY ELECTROLYSIS. 

so that to melt it without oxidation one has to be careful. 
A good way to do, experimentally, is to put the still moist 
slime in a crucible, and cover the crucible while melting. 
Even then, the escaping steam is liable to oxidize some of 
the slime, especially the antimony. Whether steam actually 
does oxidize finely divided antimony with production of 
hydrogen, has not been definitely determined. Some slime, 
especially slime which is rather dense, from anodes rather 
lower in lead, containing say 10% antimony, melts with the 
formation of little or no slag to a clean alloy. 

As lead is rather objectionable in alloys that are to be 
treated with solutions containing sulphuric or hydrofluoric 
acids, on account of the insoluble lead sulphate or fluoride 
that forms, it is desirable to get the lead out as a slag in 
melting, if possible. This may be done quite readily with 
slime from ordinary bullion by mixing the fresh slime, pressed 
as dry as possible on the filter, with enough concentrated HC1 
to convert the lead present into lead chloride. If the slime 
contains say 12% of lead, about 12% by weight of 40% HC1 
is right. Some reducing agent, as flour, may also be added 
advantageously. On melting the moist mixture in a cruci- 
ble an alloy is produced containing silver, gold, bismuth, 
some of the copper and a large part of the antimony of the 
slime, with a slag of lead chloride and antimony oxide, and a 
scum of copper and lead sulphides. The slag and scum may 
go back to the lead blast-furnace of course, where its values 
will be covered, while the metal may be electrically or chem- 
ically refined. This melting method on one occasion failed, 
and the hydrochloric acid escaped, no lead chloride or other 
slag being formed. The slime was a heavy one from lead 
containing much antimony. 



CHEMISTRY OF SL1MK TREATMENT. 65 

As an example, a light slime containing bismuth was washed 
and a moisture determination made to get the dry weight, 
which was found to be 300 gr. One hundred grams lead chlo- 
ride and about 30 cc. concentrated HO, and a few grams tar- 
taric acid as reducing agent were added and the whole stirred 
together. The mixture was added in portions to a small cru- 
cible and heated to a red heat. A good deal of smoke from 
burning tartaric acid and arsenic came off. The products 
were 135 grams of metal and about 140 of slag, and some 
additional slag was absorbed by the crucible. About 20 
g'rams of scum, that looked like galena, remained in the 
crucible. 

On analysis the slag was found to contain 9% antimony, 
1% iron, trace of bismuth and arsenic, and 1% copper, the 
remainder being mostly lead chloride. The metal contained 
30.2% silver, a trace of lead, 2.5% copper, and 13.0% bis- 
muth. The remainder was mostly antimony, not determined 
however. 

In another experiment, 224 grams slime (dry weight) melted 
with HC1, but without lead chloride, gave 82 grams metal and 
75 grams of slag. 

A complete separation of lead is thus obtained, while all 
the bismuth goes into the metal as well as about 80% of the 
antimony. 

Fused lead chloride makes an excellent electrolyte for 
depositing metals, so the slag was electrolyzed with carbon 
electrodes at a red heat. 

There are present Sb 2 3 and PbCl 2 . As the reaction, 
Sb 2 03+3PbCl 2 = 2SbCl3 + 3PbO can only take place with loss 
of energy, it need not be considered to occur. The heats of 
formation of these compound are: 



66 LEAD REFINING BY ELECTROLYSIS. 

Sb 2 3 =166,900 cal. 
3PbCl 2 = 251,700 " 



418,600 



2SbCl 3 = 182,800 cal. 
3PbO =152,400 " 



335,200 



Therefore reduction by carbon ought to yield metallic anti- 
mony in preference to lead as the reaction, Sb 2 O 3 + 3C = 2Sb + 30 
requires 79,420 cals., and the reaction Sb 2 3 + 3C + 3PbCl 2 = 
3Pb + 2SbCl 3 + 3CO requires more, namely, 146,320 cals. 

Reduction by carbon could only take place at a rather 
elevated temperature, as the reactions are stongly endothermic 
at low temperatures, and the volatility of both lead chloride 
and antimony trioxide is too great at high temperatures. 
Electrolytic reduction of the slag with carbon anode, which 
carries out the same reactions, and carbon cathode was tried 
with some success, and 18 grams of metal reduced with high 
current efficiency, containing: silver 14.5%, copper 4.7%, lead 
39.0%, antimony 40.0%. 

The presence of all this silver indicates that there were 
either metal shot in the slag to start with, or some silver was 
reduced from the scum. The quantity of silver in this pro- 
duct was relatively small, about 5.85% of the total accounted 
for. Probably the heat in the original melting was not high 
enough to thoroughly melt the slag. The temperature was 
only a red heat. 

The treatment of the alloy and similar artificial alloys 
was attempted by various methods, dry and wet. 



CHEMISTRY OF SLIME TREATMENT. 67 

For treatment with chlorine we consider the heat of 
combination of the various metals present, with chlorine, the 
figures being as follows: 

TABLE 27. 

i PbCl 2 41,950 cal. 

CuCl 35,400 " 

£ SbCl 3 30,467 " 

i BiCl 3 30,233 " 

AgCl 29,000 " 

Bismuth is capable of forming a bismuth bichloride, of 
which the heat of formation is probably somewhat greater. 
Chlorine would then apparently take out the copper, and 
then the bismuth. On passing chlorine into the metal in a 
crucible much volatile SbCl 3 came off at once, which was not 
desired or expected. 

The heat of formation of cupric chloride from cuprous 
chloride (CuCl + Cl = CuCl 2 ) is 16,000 cals., and it would 
accordingly act on the alloy as a chloridizing agent. Experi- 
mentally it could be applied more conveniently and in better 
regulated amount than chlorine. I accordingly melted to- 
gether an alloy of this composition: 

Antimony 65.2% 

Copper 13.2% 

Bismuth 21.6% 

This was put in a porcelain crucible with some PbCl 2 and 
NaCl for a cover, and 55.5% of anhydrous CuCl 2 by weight 
added, or enough to chloridize the copper of the alloy to CuCl 
and the bismuth to BiCl 2 . 

No such reactions took place. Antimony chloride came 
off in large quantity. The resulting alloy contained 31.5% 



68 LEAD REFINING BY ELECTROLYSIS. 

copper and the slag 38%. The metal also contained 19-20% 
bismuth. 

In general it has been found that precipitation of one 
metal by another from a fused melt is greatly influenced by 
the formation of compounds among the metals of the alloy 
themselves, and that the reactions are rarely complete and 
do not always proceed as indicated by the formation — heat 
figures. 

Dry chlorination of slime. — The metals can, however, be 
separated by converting them into chlorides and fraction- 
ally distilling the chlorides. 

TABLE 28. 

Arsenic chloride AsCl 3 boils at 134° C. 

Antimony " SbCl 3 " " 223° C. 

Bismuth " BiCl 3 " " 435° C. 

Copper " CuCl " " 1000° C. 

Lead " PbCl 8 " " white heat. 

For conversion into chlorides there is, of course, no 
necessity of first melting to an alloy, as the chlorine may be 
passed into the slime. 

In one experiment 250 grams of slime, containing about 

Copper 12.5% 

Bismuth 20. % 

Arsenic 15 . 7% 

Antimony 11 . 3% 

Silver 18.3% 

Lead 10 % 

was treated in a flask with chlorine. 133 grams of chloride 
distilled over, but the chlorine did not penetrate the mix- 
ture thoroughly. There is no difficulty about removing 



CHEMISTRY OF SLIME TREATMENT. 69 

arsenic and antimony as chlorides, in this way, leaving lead, 
copper, and bismuth chlorides in the residue, and also in 
distilling off the bismuth if desired. 

The heat generated by the reaction of cold chlorine 
on cold arsenic, for example, is sufficient to vaporize it and 
raise it to a very high temperature, but the thermochemical 
data to determine this temperature are not at hand. The 
reaction on the other metals is just as violent, so that on 
operations of any magnitude no external heating is necessary 
to drive off the arsenic, antimony and bismuth, and probably 
the temperature would go beyond the boiling-points of lead 
and copper chlorides, leaving silver and gold bullion in the 
melted state, with some slag of copper chloride, if the right 
amount of chlorine was used. This is not an entire innovation, 
for the treatment of gold with chlorine for removing silver and 
base metals has been successfully carried out for years, the 
process having been originated by Mr. F. B. Miller of the Syd- 
ney Mint, in 1867.* Clay crucibles are stated to be used, 
rendered impenetrable to the silver chloride by dipping them 
in hot concentrated borax solution before using. 

The chlorine for treating slime could be readily made by 
electrolyzing fused lead chloride, which is one of the easiest, 
if not the easiest, of all fused salts, to decompose electrolyti- 
cally. It has never been done commercially because lead 
chloride is not a raw material, but in the chlorination of slime 
the chlorides of copper, silver, antimony, and bismuth pro- 
duced are reducible by lead, giving the metals and lead 
chloride. 



page 615 



Rose's Metallurgy of Gold, page 441. Eissler's Metallurgy of Gold, 
615. 



70 LEAD REFINING BY ELECTROLYSIS. 

A difficulty is the storage of chlorine. The electrolytic 
plant should run continuously and the chlorine would be 
required intermittently. This has probably been one of the 
reasons why chlorination has not been applied more in metal- 
lurgy. It seems to me that an easy way to store chlorine 
is to condense it by means of sulphur to sulphur chloride, 
and produce chlorine therefrom by warming the sulphur chlo- 
ride. At temperatures below 30° C. the mixture saturated 
with chlorine has the composition SC1 4 . At 6° the compo- 
sition is SC1 2 , and at 139° SCI. Chlorine is very readily 
absorbed by the sulphur and its chlorides at the appropriate 
temperatures. In a paper on his chlorine smelting process 
Ashcroft * has described methods of pumping and drying 
chlorine. A special process for drying chlorine is not neces- 
sary in presence of sulphur, as sulphur chlorides decompose 
water. The chlorine vaporized from sulphur chloride will 
contain some sulphur, but this is a desirable circumstance, 
as any metallic oxides are readily converted to chlorides by 
sulphur chloride, even such oxides as those of aluminum and 
silicon being convertible in this way.j 

The conversion of the chlorides of antimony and bismuth 
into metal is easy in the case of bismuth, because all that 
it is necessary to do is to decompose the bismuth chloride 
with melted lead. Antimony chloride boils below the melt- 
ing-point of lead and well below the melting-point of anti- 
mony, so that it would have to be passed into melted lead 
as a vapor. 

It has been proposed f to treat slime with chlorine in the 



* Electrochem. and Metall. Industry, Vol. IV, 1906, page 96. 
fU. c. patent, Betts, 712640, November 4, 1902. 



CHEMISTRY OF SLIME TREATMENT. 71 

presence of water, producing a solution containing antimony, 
arsenic, and bismuth trichlorides, and a precipitate of lead 
chloride and cuprous chloride, insufficient chlorine being 
used to chlorinate the silver and gold (which is possible on 
account of the lower combining heat of silver, and especially 
of gold for chlorine), filtering, boiling off arsenic and antimony 
chlorides and water, and taking up the residue with water 
to remove lead chloride. The distillation of antimony chlo- 
ride solution is not as satisfactory as might be believed from 
reading descriptions of it, because it decomposes into oxide 
and hydrochloric acid, unless a large excess of HC1 is used. 
It would be difficult to find materials for carrying out the 
distillation on a large scale. The dry chlorination is then 
much superior, in which case the heat of reaction will be suffi- 
cient for the distillation, so that the apparatus question is 
not a difficult one at all in that case. 

Direct fusion with soda. — Most slime, and slime from 
ordinary grades of bullion, if dried and warmed, is very apt 
to oxidize so rapidly as to sinter or turn yellow, according 
to its composition. The oxidation takes place in two stages; 
one is a slow oxidation at a low temperature, the product 
being black and soft. If the temperature is high enough, 
the slime oxidizes rapidly, and if it contains considerable anti- 
mony, say 40 or 50%, it is not so apt to sinter, but yields a 
yellow product. This is mainly antimony pentoxide, and 
it is a difficult material to treat. It is insoluble in acids and 
infusible. It fluxes with soda, but only at a high tempera- 
ture. Heated with powdered charcoal, however, it may be 
reduced to the very easily fusible trioxide. The same end 
may be accomplished by heating it with raw, unoxidized 
slime. 



72 LEAD REFINING BY ELECTROLYSIS. 

At Trail the slime contains about 30% antimony, 20% 
silver, 10% copper, 6% arsenic, 10% lead, beside gold. 

The process worked out by the Canadian Smelting Works 
and in use there still has been described as follows: It was 
originally intended to boil the slime with sodium hydrate 
and carbonate to dissolve out the antimony,* oxidation being 
performed by drawing a current of air through the solution 
and melting the remainder in a magnesia-lined reverberatory 
to a dore bullion. The antimony failed to dissolve in more 
than very small quantities, so this step was omitted from 
the process, and the slime melted directly. Copper is diffi- 
cult to remove in this way, and this was got around to 
some extent by skimming all the dross possible from the 
lead before making anodes. 

The slime is placed in iron wheelbarrows or trucks and 
wheeled into a large brick oven, with thin walls, which can 
be heated evenly with coal fired outside. After the slime 
is pretty well dried, it is clumped into a brick stall, where 
there is a good draught, when it ignites and roasts, copious 
fumes of arsenic coming off. After oxidizing it is melted 
down in a reverberatory with soda to a dore bullion. The 
slag averages about 30-40% Sb, 5-8% Cu, 10-15% Pb, with 
considerable silica, and from 200 to 600 ozs. of silver per ton. 
The Trail refinery is using this process temporarily, until 
they have completed their experiments to devise a better 
process. To get the right amount of oxidation, which varies 
with unavoidable variations in composition and roasting of 
the slime, either coal dust or nitre, as the case may be, is 
added to the melt in the furnace. The melting part of the 

* Mines and Minerals, Vol. 25 (1905), page 28. 



CHEMISTRY OF SLIME TREATMENT. 73 

process is not satisfactory on account of the high tempera- 
ture and metal losses, nor are by-products (except copper) 
recovered in marketable form. 

In this process the antimony is probably slagged off partly 
as Sb 2 5 combined with some soda, and as Sb 2 3 . 

Melting without fluxes, slagging antimony as Sb 2 3 . — 
Antimony trioxide melts below a red heat, but contact with 
air or oxidizing gases makes the melted trioxide soon get 
pasty and finally infusible. This is because the reaction 
Sb 2 O3 + 20 = Sb 2 O5 is quite vigorous. If powdered coal or 
ground antimony be stirred in the product, another reaction 
takes place, and the easily fusible trioxide results again. 

Sb 2 3 + 20 = Sb 2 5 + 64,300 cal. at 0°, 
Sb 2 5 + 2C =Sb 2 O 3 + 2CO-5980 cal. at 0°. 

At 700° the latter reaction is still slightly endothermic, 
but it occurs readily enough. Perhaps some of the reduc- 
tion, in the case of carbon, is done by the CO produced, with 
evolution of heat. Unoxidized slime will perform the reduc- 
tion as well. 

By melting slime which is only partially oxidized, or 
proper mixtures of thoroughly oxidized and unoxidized slime, 
and keeping any excess of oxidizing gases carefully away 
during the melting, a black, glassy, and extremely fusible 
slag results, and a metallic product containing the silver and 
variable amounts of lead, antimony, and copper, according 
to the proportion of oxygen present. As slime contains usu- 
ally sulphur, a matte containing 20-30% silver and about 
50% copper is also produced, if there is not too much oxygen 



74 



LEAD REFINING BY ELECTROLYSIS. 



present. The difficulties are in getting just the right pro- 
portion of oxygen and loss of antimony trioxide by volatili- 
zation. At the temperature necessary to melt the silver 
antimony trioxide volatilizes very fast. If some antimony 
and lead are left in the silver by deficiency of oxygen, the 
temperature may be much reduced, but the metal requires 
another treatment to remove antimony and lead. If the 
melt is not well covered, the pentoxide will form, so that a 
reverberatory furnace is not suitable, both for this reason 
and on account of the volatilization difficulty. No satisfac- 
tory crucible has been found, as the slag attacks most cru- 
cibles rapidly. My experiments indicate, however, that a 
cast-iron crucible will do quite well. The electric furnace is 
the remaining means, and is entirely feasible from a power 
standpoint and admits of melting large quantities rapidly 
with little loss. The specific heat of slime can be roughly 
•calculated as follows: 





TABLE 29. 


Heat in melted silver 


at 960° 


Heat in Sb 2 3 


" 960° 


Heat in Pb 


" 960° 


Heat in Cu 2 S 


" 960° 


Volatilization AS 2 3 


" 960° 



Heating and vaporizing H 2 ' ' 960 c 



at 960° per lb. 89. 15 lb. cals. 
200 
150 
200 
200 
700 



These are the principal constituents. The figures are 
not known for antimony, lead, and arsenic oxides, and can 
only be very roughly got from comparison with compounds 
for which the exact figures are known. The above figures 
will do for our present purpose, as will be readily seen below. 

Suppose the slime contains «. 



CHEMISTRY OF SLIME TREATMENT. 75 



TABLE 30. 

H 2 15% = 102.5 lb. cals. 
Silver 30%= 27.0 
Sb 2 3 30%= 60.0 
As A 10%= 20.0 

PbO 10%= 15.0 

Cu 2 S 5% = 10 

Heat required in lb. cals. 234.5 



(The pound calorie is the amount of heat required to 
heat one pound of water, one degree centigrade, and is equiv- 
alent to .00052 K.W. hours, or .00069 E.H.P. hours, in elec- 
tric energy.) 

The heat necessaiy at 100% efficiency per pound of slime 
is 0.12 K.W. hours; at 50% efficiency, which is easily obtained, 
and for lead producing 80 lbs. of slime per ton, 20 K.W. hours 
would be necessary for melting. This is so small that 
quite large proportionate errors in the specific heats above 
would not make any practical difference. 

Xot all types of electric furnace would be suitable. Con- 
tact with hot carbon would tend to reduce the slag and render 
the precipitated metal too base. A furnace heated by radi- 
ation from an arc or a heated carbon rod would do, but a 
furnace of the resistance type, in which the heat is gener- 
ated in a narrow conductor of the metal, would seem to be 
best adapted. The heating current may either be induced 
as in the Colby or Kjellin * and similar furnaces. That 
electric furnaces will be used in slime and silver melting is 
probable. Small induction furnaces for melting steel and 



* Colby, U. S. patents 428378 and 428379, May 20, 1890; Gin., 
771872, Oct. 11, 1904; Schneider, 761920, June 7, 1904; Betts, U. S. 
Patent 816558, April 3, 1905. 



76 LEAD REFINING BY ELECTROLYSIS. 

brass are now on the market and in successful use,* but the 
material of which the crucibles are made would have to be 
changed for slime melting, and some arrangement would be 
necessary to catch the fumes given off. 

When the slime contains bismuth in appreciable quantity 
the melting process is at its best, because the bismuth is in- 
termediate in oxidizability between silver on the one hand, 
and lead and antimony on the other, and as a consequence 
the percentage of oxidation does not need to be so carefully 
controlled to insure a separation of gold and silver from lead 
and antimony. In case the oxygen is higher than usual,, 
more bismuth is slagged; in case the oxygen is lower, more 
bismuth goes into the dore, while in either case the antimony 
and lead remain almost entirely in the slag, and the silver as 
metal. Also the presence of bismuth in the silver increases 
its fusibility, so that the melting temperature need not be 
nearly so high. The slag high in Sb 2 3 is so fusible that it 
melts below a red heat. 

These facts are brought out well from the analysis of the 
products resulting from the melting down of some partially 
air-oxidized slime in my laboratory. 

TABLE 31. 

Metal 35 Gr. Slag 80 Gr. Matte about 2 Gr. 

Au 78% 

Ag 66.23% 

Bi 20.3% 2.95% 30% 

Cu 5.1% 1.15% 46.3% 

Sb 1.3% 28% None. 

Pb .8% 34.9% None. 

The value, 5% for copper in metal, is probably too high, 
as the sample may have contained a little intermixed matte. 

* Electrochemical and Metallurgical Industry, 232, 1907. 



CHEMISTRY OF SLIME TREATMENT. 77 

On cooling the bar bismuth liquates out in little drops 
that can be knocked off. These contained little beside bis- 
muth. The analysis showed 87% bismuth, 6% silver. 

The treatment of these products can be carried out suc- 
cessfully. The bullion may be parted by the methyl-sulphate 
method, as described in Chapter IV, and the slag may be freed 
from bismuth, if required, by melting it with a little anti- 
mony. The residual slag yields its antimony to hydrofluoric 
acid, from which solution it may be deposited electrolytically 
as described in Chapter III. About one-half the copper pres- 
ent also dissolves in the HF, the removal of which will be 
taken up with the description of antimony depositing. 

Dilute nitric acid was tried for dissolving lead and bis- 
muth from the slag, after which the residue could be con- 
verted into antimony or antimony compounds. If the nitric 
acid is not very strong little or none of the antimony is con- 
verted to higher oxides, as it is difficult to peroxidize it. 
Nitric acid acts slowly on the slag, finally leaving a soft light 
yellow, rather dense residue of antimony oxide and a solu- 
tion of lead nitrate, which can be easily crystallized. 

Mattes high in silver are analogous to the sulphides made 
from silver ores in a hyposulphite leaching mill. There are 
several methods for treating such material. Stetefeldt's 
process of melting the sulphides in an iron pot, roasting, and 
dissolving out copper sulphate with water in presence of 
metallic copper to precipitate any silver in solution, would 
seem to be applicable to the present material from the roast- 
ing stage on. The residue from the copper extraction con- 
sists mainly of silver. The roasting was, as described by 
Stetefeldt, performed in a small muffle-furnace after being 
first ground in a small ball-mill. Details will be found in 



78 LEAD REFINING BY ELECTROLYSIS. 

Stetefeldt's, " The Lixiviation of Silver Ores with Hyposul- 
phite Solutions," and Collins' "Metallurgy of Silver." 

The sulphuric acid process (Dewey- Walter process) con- 
sists in boiling the sulphides with hot concentrated sulphuric 
acid in an iron pot. The sulphuric acid oxidizes the sulphur 
of the sulphides as well as the metals. It seems reasonable 
to suppose that the matte, if ground fine, would react simi- 
larly to precipitated sulphides. If so, this method would be 
much simpler and cheaper than the roasting method. Details 
will be found in Collins' " Metallurgy of Silver." 

I have found that these mattes are as readily converted 
to bullion by the following process: Grind the matte, and 
add in an iron pot enough concentrated sulphuric acid to 
react with all the silver and copper as follows: 

Ag 2 S + 2H 2 S0 4 = 2Ag + 3S0 2 + 2H 2 0, 
Cu 2 S + 2H 2 S0 4 = 2Cu + 3S0 2 + 2H 2 0. 

On heating the mixture gently part of the matte is con- 
verted to sulphates, and a little sulphur comes off with much 
S0 2 . The dry product is transferred to a melting crucible, 
and treated, when the mass melts down quietly, to copper- 
silver bullion. 

Melting with the addition of sulphur for matte and slag. — 
In case the slime contains little bismuth, or only small quan- 
tities of silver, the direct fusion of the partially-oxidized slime 
suffers from two disadvantages. One is the high tempera- 
ture necessary to melt the dore bullion, which is too high 
for antimony trioxide to remain as liquid, and on the other 
hand, either the silver is apt to go into the slag, or the dore 
contains too much Lead and antimony. 



CHEMISTRY OF SLIME TREATMENT. 79 

A very neat melting method consists in adding sulphur 
to the air-oxidized slime, in about sufficient quantity to 
reduce any Sb 2 5 to Sb 2 3 and to form a silver-copper matte 
with the copper and silver of the slime.* 

As an example Trail slime containing, when dried, 

Ag 14.6% 

Cu 8.1% 

Sb 27.60% 

Pb. . 16.0% 

As 27.0% 

Au 34 ozs. per ton. 

was melted with various amounts of sulphur from 8 to 12%. 
8% was found to be about right. When the slime was given 
a slight further roast as a preliminary, a little more sulphur 
was used. 

100 grams roasted slime and 10 grams sulphur, melted in 
a porcelain crucible, gave 

TABLE 32. 

Matte. 43 Gr. Slag, 45 Gr. 

Ag 34.0% Cu 0.2% 

Cu 19.2% Pb 12.1% 

Pb 24.8% Sb 51.4% 

S 13.9% As 5.2% 

Au 25% Fe 3.0% 

Sb 5.8% 

Other mattes from the same slime contained: 

TABLE 33. 

Ag 41.7% 38.1% 46.8% 

Cu 23.2% 21.8% 26.3% 

Pb 17.5% 5.8% 

* Patent applied for. 



"80 LEAD REFINING BY ELECTROLYSIS. 

100 grams of the unroasted but thoroughly oxidized slime 
gave with 7 grams of sulphur 35 grams of matte and 46 grams 
of slag. This matte contained less lead (about 5%) and. less 
antimony. 

The melting temperature was low, about 600° C. The 
matte should have no action on iron, and the slag might not 
either, so I tried a cast-iron pot for melting. 

400 grams of slime and 28 grams of sulphur were added 
to the red-hot pot, and melted clown. Large quantities of 
arsenic came off. The product consisted of 140 grams of 
matte with considerable metal intermixed, and about 190 
grams of slag. 

Another melt with 32 grams sulphur gave a product with 
no metal, but the cast was spoiled, and the products melted 
over again, with the addition of a few grams of sulphur. 
Products were about 160 grams matte and 165 grams slag. 

After these three melts the pot shows no sign of wear. 
The last slag, however, analyzed 8.25% iron as against 3% 
when melted without access of iron. On the other hand, 
the fusion in an iron crucible gave a slag lower in silica, as of 
course would be expected. 

Treatment of the slag: (1) This has been reduced to 
hard lead by smelting with litharge and carbon, from which 
first the lead may be extracted electrolytically, see page 
5:, and the antimony residue refined with the fluoride so- 
lution. 

(2) It can be leached, after grinding, with hydrofluoric 
acid for antimony fluoride solution. The action of hydro- 
fluoric acid is rapid and evolves considerable heat. The anti- 
mony can be deposited from the fluoride solution, to which 
sulphuric acid should be added. If the hydrofluoric acid 



CHEMISTRY OF SLIME TREATMENT. 81 

used contain also sulphuric acid, the residue will consist 
mostly of lead sulphate. 

(3) The slag may also be leached after grinding wiih 
dilute nitric acid. After several hours' action the residue 
consists of yellow antimony trioxide and the solution con- 
tains lead nitrate. The production of lead nitrate as a by- 
product of a lead refinery is analagous to the production of 
copper sulphate by a copper refinery. 

(4) The slag can be reduced to metal electrolytically, 
either with a fused electrolyte as lead chloride, see page 66, 
or with an aqueous electrolyte, as sulphuric acid solution. 
The following experiment illustrates the latter: . 

100 grams of slag containing about 50% antimony, 8% 
iron, 5% arsenic, and 15% lead, were ground rather fine, say 
30 mesh, and placed in a shallow lead pan about 3 by 4 inches 
and about 1 inch deep. 25% sulphuric acid was added and 
a horizontal lead anode used about 2} by 3^ inches, wrapped 
in cloth and about \ inch from the slag. The current varied 
from three to six amperes, equivalent to a current density 
40-80 amperes per square foot. During most of the run 
no gas was seen to rise from the cathode. About the middle 
of the run red antimony sulphide appeared in the electrolyte 
and floated around, the total quantity produced being 2.45 
grams. Current necessary for complete reduction about 
38 ampere hours, and when 36 ampere hours had passed H 2 S 
came off for a while, and then was replaced by hydrogen. 
The appearance of H 2 S would be a good indication of the end 
of the reaction. 

Any iron of course went into solution as ferrous sulphate. 
It was thought that fluorine and silica might both be present 
in the slag and give rise to the formation of H 2 SiF 6 during 



82 



LEAD REFINING BY ELECTROLYSIS. 



the action. Acids forming soluble lead salts cause a rapid 
corrosion of lead anodes in sulphuric acid, so I rather expected 
lead sulphate would drop off the anodes into the slag during 
the run, and used a cloth to keep any from dropping into the 
slag. The lead anode, however, was not attacked. 

The cathode area was 10-12 square inches; anode area, 
8.9 square inches; current 3-6 amperes, e.m.f. 2.25 to 2.9 
volts. 

TABLE 34. 



Time, 
Hrs. Min. 


Volts. 


Amperes. 


Amperes, 
Hours. 


Remarks. 


0.15 


2.25 


4.3 




No gas noticed coming from cathode. 


0.30 


2.25 


3.5 




No gas noticed coming from cathode. 


2.30 


2.25 


3.5 




No gas noticed coming from cathode. 


2.45 


2.45 


3.0 


*9^85 


No gas noticed coming from cathode. 


5.35 


2.50 


3.5 


31.67 


No gas noticed coming from cathode. 


9.28 


2.60 


3.0 


31.25 


Sb 2 S 3 suspended in electrolyte. 


9.58 


2.60 


3.2 




Sb 2 S 3 suspended in electrolyte. 


10.45 


2.6 


3.2 




Some H 2 S evolved. 


11.35 


2.75 


4.6 






12.20 


2.65 


5.8 






12.45 


2.65 


5.0 


44.93 




13.45 


2.9 


5 






15.05 


2.7 


3.6 




No H £ S evolved. 


16.25 


2.8 


3.5 






17.05 






55 





The rise in voltage after about 12 hours and 45 minutes 
indicated the completion of the reaction. 

The electrolyte contained some ferrous sulphate. Part of 
the former slag was cemented to the lead tray and part was 
loose. It had not changed much in appearance. 

After washing, and before drying, part of the product 
was rapidly melted in a crucible, producing metal and no slag, 
showing good reduction. Another test was made and gave 
15.5 gr. metal and 4.5 gr. slag. In the first test a clay cruci- 



CHEMISTRY OF SLIME TREATMENT. 83 

ble was used which probably had absorbed the small quan- 
tity of slag. 

To reduce the 40-45 grams antimony and 15 grams lead 
in the slag would require about 40 ampere hours, which it 
will be noticed corresponds with the increase in voltage. The 
current efficiency appears from the slight evolution of hydrogen 
ami from the above noted circumstances, to be quite high. 
The slag produced in melting may have resulted from oxi- 
dation during drying and fusion. 

These slags may also be conveniently reduced to hard 
lead, by smelting with litharge or dross from the refined lead 
pot and carbon. 

The extraction of the lead from the hard lead is achieved 
by refining with the fluosilicate electrolyte. Further experi- 
ments are necessary before it is possible to say whether it 
would be better to melt the antimony residue and refine elec- 
trolytically with the fluoride solution, or refine the antimony 
skeleton directly as anode in the same solution. 

Direct electrolysis with slime as anode. — As the slime 
retains very often the form of the original anode, especially 
when the percentage of impurity is rather high, say 3 to 10%, 
the proposition of placing the anode with the slime still adher- 
ing in an appropriate solution, and making it the anode, 
while one of the principal metals contained goes over to the 
cathode, seems promising. 

Some means of removing or washing out the lead electro- 
lyte in the slime, worth on the average, say S2.25 per ton 
of lead refined, is very necessary. In time, with frequent 
changes of wash-water in a tank containing a set of anodes, 
this solution could be removed to any desired extent by the 
simple action of diffusion. AVe tried another method, namely, 



84 



LEAD REFINING BY ELECTROLYSIS. 



connecting up the anode with attached slime as cathode in 
a solution of fluosilic acid, with carbon anodes. Passing the 
current deposits some lead in and on the slime, and lead 
peroxide on the anode, while the valuable SiF 6 present of 
course moves away from the slime, now cathode, into the 
solution. A removal is possible in this way, but our result 
from the following experiment could not be called successful. 

The original lead alloy contained lead 90%; arsenic 5%; 
antimony 3.0%; copper 1%; silver 0.7%, and Bi 0.05%, and 
was refined with an electrolyte containing 8.05 grams lead 
and 15.9 grams SiF 6 per 100 cc. The anode was originally 
about J" thick, and was treated until the lead was practi- 
cally all removed. 

The residue was then made cathode in a solution contain- 
ing 6.15 grams SiF 6 per 100 cc. and no lead. Volume of solu- 
tion 715 cc. = 46.5 gr. SiF 6 . Anodes of carbon (f" round 
electric light carbons). 

TABLE 35. 



Time. 


Amperes. 


Volts. 


Amperes 
per Square Foot. 


Temperature. 


9.30 

9.45 

10.20 

11.30 

4.00 


1.82 
1.82 
1.45 
1.93 
1.45 


2.45 
2.40 
2.25 
2.40 
2.40 


12.5 

12.5 

9.9 

13.2 

9.9 


16° C. 

17° C. 
19° C. 

20° C. 



The SiF 6 in the solution increased by only 8.5 grams 
instead of 10.7 grams or more present in the slime. 

The volume of the pores in the slime treated was 67 cc. 
While sufficient current was passed to remove 29.1 SiF 6 grams 
at 100% efficiency, only part of that present was actually 
removed, with an efficiency of about 30%. 



CHEMISTRY OF SLIME TREATMENT. 



85 



It is possible to so choose the electrolyte for refining the 
attached slime, that the fluosilicic acid and lead fluosilicate 
of the slime is not lost, but may be recovered from the second 
electrolyte. 



1.75 



1.50 



1.25 



1.00 



.75 



.50 



.25 















/ 












/ x 


/ 
























*/ 






















: 


c^-^^ 























9 10 

Time Hours 
Fig. 10. 



11 



12 



Experiment. — Bullion containing arsenic 5%, antimony 
3%, and copper 1% in the shape of an anode, 3JXliX* 
^§ inches, was electrolyzed until 300 grams lead was dissolved 
out, or about 350 grams of alloy decomposed. The core was 
then about §" thick. The slime was scraped off one side for 



86 LEAD REFINING BY ELECTROLYSIS. 

another experiment and the other half made anode in a hot 
solution containing 15% CuS0 4 -5H 2 and H 2 S0 4 . Tem- 
perature about 70° C. On the start with a current of 1 ampere 
the voltage was .26, rising after about six hours as plotted 
in Fig. 10. The copper deposited remained good, though 
the voltage finally increased very much. A small quantity 
of Sb 2 3 was found in the solution, but nearly all remained 
in the slime, with some lead sulphate. 

Of course the concentration of copper in the solution 
continually got smaller, for little was supplied by the anode 
to make up for that deposited at the cathode, the result of 
the electrolysis being expressed by these reactions: 

3CuS0 4 + 2Sb = 3Cu + Sb 2 3 + 3H 2 S0 4 ; 
3CuS0 4 + 2As = 3Cu + As 2 3 + 3H 2 S0 4 ; 
CuS0 4 +2Ag= Cu+Ag 2 S0 4 . 

The peculiar flat place in the curve corresponds in volt- 
age to the formation of silver sulphate in the slime, and may 
be due to silver dissolving at that time. 

Twelve grams of copper were deposited, about 9.5 before 
voltage reached A volt and 2.5 grams thereafter. Roughly 
there were in the slime 7.5 grams arsenic, 6 grams antimony, 
and 1.5 grams copper, equivalent to the deposition of 16.0 
grams of copper, so that the oxidizing efficiency at the anode 
was about 75%. 

The anode, with the slime still remaining on it, was put 
in dilute HF + H 2 S0 4 , with the idea that the antimony would 
dissolve out and the slime drop off, which would be a good 
thing in practice as it would save cleaning the anode scrap. 
However, the slime did not fall off, perhaps because the HF 



CHEMISTRY OF SLIME TREATMENT. 87 

used was too small in quantity, theoretically only just about 
sufficient being used to form SbF 3 , while an excess should have 
been used on account of the long time necessary to complete 
the reaction. 

To make use of this simple process the following condi- 
tions are necessary: 

(1) Anodes containing enough impurity to cause the slime 
to hold together quite firmly. (2) Rather thin anodes, in 
order that the electrolytic action can penetrate all the slime. 
(3) A supply of copper sulphate, or copper containing mate- 
rial, as matte, which could be treated with the dilute sul- 
phuric acid produced in the process to make copper sulphate 
solution. (4) Recovery of H 2 SiF 6 from the copper sulphate 
sulphuric acid solution, as by precipitation with K 2 S0 4 , and 
distillation of the sodium fluosilicate with sulphuric acid to 
recover H 2 SiF 6 . 

The arsenic of course dissolves in the hot solution and 
can be crystallized out merely by cooling. Analysis of the 
figures noted in the experiment indicate that the electrolysis 
goes smoothly, as long as metallic arsenic remains in the 
slime, and thereafter the voltage rises. 

The recovery of antimony and hydrofluoric acid wotild 
be by the usual method described in Chapter III. 

Such a slime process can be carried out with other solu- 
tions than that of copper sulphate and sulphuric acid; for 
example, antimony or copper fluoride-hydrofluoric acid solu- 
tion which possesses the advantage of taking nearly the 
entire slime; that is, the arsenic, antimony, and copper into 
solution, while depositing copper at the cathode. 

In refining bullion containing Pb 65.37%, Bi 7.32%, Sb 
19.51%, As 5.85%, and Ag 1.95%, an anode covered with 



88 



LEAD REFINING BY ELECTROLYSIS. 



slime was electrolyzed in a solution of SbF 3 + HF + Na 2 S0 4 , 
containing about 4% antimony. The cathode deposit was 
black and spongy and the e.m.f. very high, so that it was 
not a success. This may be due to the fact that t'he solution 
contained SO/', which I afterward found out was a mis- 
take. When SO4" is present the antimony is converted at 
the anode partly into insoluble basic insulating compounds, 
while if the only anion present in W, the antimony goes 
straight into solution as SbF 3 . 

An alloy containing Pb 65.56%, Ag 1.94%, Cu 1.94%, Bi 
6.84%, As 5.47%, and Sb 18.24%, after removing the lead 
electrolytically, was treated in the same way, only all PbSiF 6 
was removed by steeping the anode, after extracting lead, 
in hot water. The antimony electrolyte contained 6.4 grams 
Sb, as SbF 3 , 5.4 grams HF and 5 grams (NH 4 )2S0 4 per 100 cc. 

The results are given in Table 36. 

TABLE 36. 











Back 




Time. 


Amperes. 


Volts. 


A 

Amperes, 
Sq. Ft. 


E.M.F. 
Volts. 


Remarks. 


0.5 hours 


0.36 


0.42 


6.7 


0.23 




5.0 " 


0.38 


0.38 


6.1 


0.14 




8.0 " 


0.34 


0.38 


5.4 


0.14 


Fair deposit. 


14.1 « 


0.29 


0.34 


4.6 


0.15 




17.i " 


0.33 


0.38 


5.3 


0.18 




17.i " 


0.33 


0.36 


5.3 


0.18 




26. | " 


0.28 


0.38 


4.5 


0.18 


Fair deposit. 


136 


0.10 


1.24 


1.6 




Powdery deposit. 



The slime was irregularly attacked, being very hard in 
places and very soft in others. The core had been attacked 
seriously under the soft spots, showing the current was 
applied for too long a time. 



CHEMISTRY OF SLIME TREATMENT. 89 

Fusion to alloys and electrolytic refining. — In the para- 
graph on "fusion", page 04, a description is given of an 
easy method of eliminating lead from the alloy if desired. 
It is then possible to eliminate lead in the first place, so if 
the presence of lead is objectionable on account of any elec- 
trolytic process desired to be used, it may be disposed of on 
the start. 

The treatment of the alloy by electrolysis will of course 
depend on the composition of the alloy. If it should be 
mostly bismuth, as it might be in working up bismuth alloys, 
it could be electrolytic ally refined with a solution contain- 
ing about 10% free methylsulphuric acid and 4% of bismuth 
as methylsulphate. This makes an excellent electrolyte for 
refining bismuth * If the alloy is mainly silver, copper, or 
antimony, refining with the solution appropriate for that 
metal could be adopted. 

For the usual case, producing a mixed alloy with rela- 
tively large amounts of silver and antimony and important 
amounts of lead, copper and perhaps bismuth, this method 
does not seem to offer sufficient advantages over the usual 
wet methods of treatment. In this case it would be advisable 
to add outside antimony, silver, or copper that would be 
refined anyway, to make a preponderating proportion of one 
metal, which is always desirable in electrolytic refining. Of 
the three metals copper offers the chief advantages. Copper 
for refining is usually available, it produces the most satis- 
factory cathode deposit, and comes down as pure metal, 
and most important of all the resulting slime was found to 

* According to Mohn, bismuth trichloride, with 10% free HC1, is used 
successfully for this purpose. Electrochemical and Metallurgical Industry, 
August, 1907. 



90 LEAD REFINING BY ELECTROLYSIS. 

contain only traces of copper or antimony, and thus a sharp 
separation and complete recovery of the metals is easy. 
There is no difficulty in depositing pure copper in presence 
of antimony fluoride, even though the percentage of copper 
falls to a very low figure, if a brisk circulation is kept up. 
This then affords a separation, copper on the cathodes, anti- 
mony collects on the solution and continually increases in 
amount while copper diminishes, and silver, bismuth, and 
gold constitute the slime. 

In one experiment the alloy, which contained approxi- 
mately 30% silver, 45% antimony, 14% bismuth, and 9% 
copper, was melted in a crucible and two and one-half times 
its weight of copper added, under a cover of salt, and then 
contained 8.6% silver, 12.9% antimony, 4% bismuth, 74% 
copper. There was no volatilization during the process, and 
the alloy was very much more readily fusible than copper. 
The maximum temperature during the melt did not exceed 
800° C. probably. 

The alloy was electrolytically refined with a solution 
containing CuF 2 , NaF, and HF, and no H 2 S0 4 . Copper 
was deposited in a satisfactory condition on the cathodes, 
though it was not as solid as the copper from a sulphate solu- 
tion. The cathodes were bright and clean looking, but the 
crystallization was coarser than is obtained with a sulphate 
solution. 

Some white bismuth compound separated on the bottom 
of the tank during electrolysis, and the anode slime contained 
the remainder of the bismuth, as well as the silver, and 1% 
copper and 2.5% of antimony. The extraction of antimony 
and copper was then approximately 96% and 99.7% of that 
present respectively. The slime would be melted for dore 



CHEMISTRY OF SLIME TREATMENT. 91 

bullion and a bismuth slag, and the solution, which continu- 
ally diminishes in copper content while antimony increases, 
would have to be treated for metallic antimony and regener- 
ation of the copper fluoride; for example, by precipitating the 
remaining copper by running the solution through broken 
antimony, depositing antimony from the solution with a 
lead anode and neutralizing HF in the solution with copper 
oxide, roasted matte, etc. 

Wet treatment with regeneration of the solutions. — By 
getting the metals of the slime into the solution in some way 
and electrolyzing the solution for the contained metals, and 
simultaneously producing an anode product which may be 
used in oxidizing further quantities of slime, important advan- 
tages may be secured. It not being necessary to dry the 
slime, the danger of loss from dutsing and the physical discom- 
fort of working with such poisonous, penetrating dust, are 
avoided, as well as the involved expense. For any w T et treat- 
ment process the raw T slime is better suited than it is after 
drying on account of its more open nature. The chemicals 
required, an imporant item in most methods, are reduced 
in amount, only enough being required to make up for 
mechanical losses. A certain amount of electric energy is 
needed, but this is not great. It will be noticed that in 
most of the other methods outlined, steps are introduced 
which require electric energy, so this item applies to prac- 
tically all methods anyway, and is not therefore a disadvant- 
age peculiar to this class of processes. 

The acids, the salts of which have been used in the solu- 
tions, are few in number, the choice being necessarily limited. 
The ideal acid to use is the same as that used in the lead 
depositing electrolyte, that is, fluosilicic acid. A combina- 



92 LEAD REFINING BY ELECTROLYSIS. 

tion of lead peroxide as oxidizing agent and fluosilicic acid 
solution has an important advantage on the score of sim- 
plicity of the whole. With their use it would not be neces- 
sary even to wash the slime, as the two electrolytes, slime 
treating and lead depositing, are the same, and after suitable 
purification of the slime electrolyte can be exchanged when 
convenient. 

With the fluosilicate electrolyte we may use lead per- 
oxide as oxygen-carrier, and the other metal that has been 
used as oxygen carriers from the anode to the slime is iron, 
passing from the ferrous to the ferric state, and back again. 

The production of ferric fluosilicate has not been seri- 
ously attempted, and there are certain objections to its use. 
One comes from the fact that in electrolyzing the solution 
with insoluble anodes of carbon for ferric fluosilicate, silica 
deposits on the anodes and stops the oxidation of the iron, 
while if hydrofluoric acid, in small quntity, is used to pre- 
vent this, the difficulties in the way of a successful diaphragm, 
materials containing silica being barred, have prevented any 
serious attempt in this line. 

Fluosilicic acid being unsuitable, we find that sulphuric, 
hydrochloric, and hydrofluoric acids are cheap enough to be 
considered from slime-treating solutions. The question of 
tanks is an important one, on which depends to a large 
extent the choice of available acids. Until recently, there 
appeared to be no suitable tank for working with strongly 
acid chloride solutions (to keep SbCl 3 from decomposing). 
However, a concrete tank, saturated with sulphur, described 
in Chapter VII, is not acted on by hydrochloric acid. 

Ferric chloride has been used in experiments on anti- 
mony slime from refining hard lead. The electrolyzed solu- 



CHEMISTRY OF SLIME TREATMENT. 93 

tion contained beside hydrochloric ferrous chloride and anti- 
mony trichloride, which yielded antimony on the cathode 
and ferric chloride at the anode.* 

Ferric sulphate is an ideal solution in some ways, on ac- 
count of the cheapness with which it may be produced and 
the ease of handling it in lead-lined tanks. It also has an 
advantage in separating copper and arsenic in solution, from 
antimony hydroxide and silver in the undissolved portion. 

Hydrofluoric acid may also serve as a basis of the solu- 
tion, and ferric fluoride be formed at the anode as oxidizer. 
Hydrofluoric acid may also be used in connection with ferric 
sulphate, when the antimony will go into solution as anti- 
mony trifluoride. 

The use of the insoluble anode product, lead peroxide, has 
also been tried. Lead peroxide is easily obtained in quan- 
tity in dense scales and plates by electrolyzing lead fluosili- 
cate solution with a graphite anode. 

Lead peroxide and metallic lead, copper, etc., in slime 
react with lead peroxide in the presence of fluosilicic acid, 
for instance, to form fluosilicates. 

Pb + Pb0 2 + 2H 2 SiF 6 = 2PbSiF 6 + 2H 2 0. 

Ferric sulphate process. — This is a neat process, and has 
been the subject of a great deal of experimenting. It is 
applicable to slime from copper refining as well as to lead 
slime. In the treatment of copper refinery slime it is apt to 
find a large use. The description may also prove of interest 
in connection with the Siemens & Halske process for copper 
ores and matte. f 

* A. G. Betts, Trans. Am. Electrochemical Society, Vol. VIII, page 188. 
t Borcher's " Electric Smelting and Refining." 2d. Eng. Ed., page 260. 



D4 LEAD REFINING BY ELECTROLYSIS. 

Ferric sulphate is a very soluble salt, of which a syrupy 
solution containing 10% Fe" may be readily prepared. This 
is too strong for work with slime, principally on account of 
the less solubility of the resulting ferrous sulphate. Solutions 
used for slime treating should contain about five volume per 
cent of iron. 

Ferric sulphate solution reacts with slime very readily, 
oxidizing metallic copper and cuprous sulphide to copper 
sulphate, antimony to hydrated trioxide, arsenic to arsenious 
acid, bismuth to basic sulphate, finely-divided lead to lead 
sulphate, and when hot converts silver to silver sulphate. 
The complete oxidation of silver is difficult or impossible on 
account of the reducing action of the ferrous sulphate 
formed. Practically, to dissolve one-half to two-thirds of 
the silver in slime requires the use of a considerable excess 
of ferric sulphate, so that the process is simplified in some 
respects by only using enough ferric sulphate to oxidize the 
other metals. When slime contains sulphur or sulphides, 
which it almost always does, especially copper slime, the solu- 
tion of silver is hindered or entirely prevented. Apparently 
in the reaction between ferric sulphate and silver little or 
no energy is liberated and the presence of finely-divided sul- 
phur, combining with the silver, can even reverse the action 
giving 

Ag 2 S0 4 + 2FeS0 4 + S = Ag 2 S + Fe 2 (S0 4 ) 3 . 

Tellurium is dissolved by ferric sulphate and may be pre- 
cipitated out on metallic copper, as a greasy, black coating, 
Selenium and gold are not dissolved from slime by ferric sul- 
phate. 



CHEMISTRY OF SLIME TREATMENT. 95 

As the solution is prepared by electrolysis, and it is ad- 
visable to have a solution of as high conductivity as possible, 
the ferric sulphate used will contain some free sulphuric 
acid. 2-5%, and some ferrous sulphate, usually equivalent 
to 1% ferrous iron, Fe". 

From the reactions taking place, it will be seen that the 
process is not entirely cyclic. The reaction of the copper 
of the slime, Cu + Fe 2 (S0 4 )3 = CuS04 + 2FeS0 4 , is directly 
reversed in the electrolytic tank, so, as far as copper is con- 
cerned, the same solution could be used over and over again 
without any additions being made to it. In treating copper 
slime, consisting largely of Cu 2 S and silver, this condition 
is quite well realized. 

The reaction on antimony and arsenic, which consumes 
a good percentage of the ferric iron used, is not reversible. 
Oxygen is removed from the solution in the insoluble anti- 
mony hydrate, and arsenic removes oxygen from the cycle 
though not from the solution. Lead removes SO/', but not 
in large or serious quantity. 

For many slimes, containing say 30% Sb, 15% Cu, 10% As, 
beside lead and silver, the amount of iron reduced by anti- 
mony and arsenic would approximate two-thirds of the total. 

Several methods exist of adding combined oxygen to the 
solution to make up the deficiency, but the best seems to be 
the addition of copper oxide in some form, especially as roasted 
copper or copper-lead matte. As the copper is recovered as 
electrolytic metal, from a raw material, the process may be 
credited with part of the enhanced value of the copper. With 
the addition of copper oxide from any source, and crystalli- 
zation from the solution of arsenious acid, the solution may 
be used over, if mechanical losses are made up. 



96 LEAD REFINING BY ELECTROLYSIS. 

Another method of supplying oxygen consists in air-dry- 
ing the slime before treatment with ferric sulphate, which 
introduces considerable oxygen, but this suffers from several 
objections, such as the formation of hard lumps which are 
attached with difficulty and greater expense and losses. 

The separation of the metals by the sulphate solution is 
not perfect, principally because antimony and bismuth hy- 
droxides or basic salts are not entirely insoluble in the solu- 
tion. The solubility of Sb 2 3 in the solution is approximately 
1.6 grams per litre cold and 2.2 grams hot. Variations in 
the percentage of sulphuric acid have little influence on the 
solubility of antimony. The amount of bismuth dissolved is 
about 1.5 grams per litre, and none separates on cooling. 
From other results obtained, the solubility of bismuth is 
somewhat greater in the cold solution, on account of change 
to another series of so far uninvestigated salts in which 
bismuth has greater basicity. 

For the sake of example, let us assume that one ton of 
lead contains 7.4 lbs. silver, 2 lbs. bismuth, 4 lbs. arsenic, 
10 lbs. copper, 20 lbs. antimony, while 5 lbs. of lead will also 
remain in the slime. 

Ferric iron required is readily calculated by the use of 
factors, as given in Table 37. 





TABLE 


37. 






Foi 


• silver 






None. 




" 


bismuth 


2X 


.81 = 


1.62 


lbs. 


1 c 


arsenic 


1X2 


.24 = 


8.96 


1 1 


1 1 


copper 


10X1 


.76 = 


17.60 


'•' 


1 1 


antimony 20X1 


.40 = 


28.00 


" 


( e 


lead 
Total. . 


5X 


.54 = 


2.7 

58.88 


! 1 



CHEMISTRY OF SLIME TREATMENT. 



97 



This amount of ferric iron is contained in about 23 cubic 
feet of electrolyzed solution. Allowance must be made for 
the fact that the action is not entirely complete. Usually 
not all the copper and arsenic present dissolve, but only 
about 90%. Taking account of the solubility of bismuth and 
antimony, and of copper already in solution to the amount of 
10 grams per litre, and used to boil out traces of silver and 
reduce excess of Fe", the distribution of the products is 
about as follows: 



TABLE 38. 



In Solution. 



In Sediment After Cooling. 



In Residue. 



All Iron. 
2.35 lbs. antimony 
1.8 " bismuth 
26 . " copper 
8.1 " arsenic 



Lbs. 
85 antimony 



All Lead. 

16.8 lbs. antimony 

.2 " bismuth 

1 ' ' copper. 

.9 ' ' arsenic. 



The solution also contains fluosilicic acid present in the 
slime on account of not entirely complete washing. This 
is a troublesome compound to have present, as will be ex- 
plained elsewhere, on page 109. 

The sediment can be put in a charge of fresh slime. The 
residue has then to be washed rather free of iron and copper 
sulphate, and is then treated with a solution of hydrofluoric 
and sulphuric acids, which may vary largely in composition, 
about 5% sulphuric acid and 5-10% hydrofluoric acid being 
satisfactory. The result is the solution of approximately 
95% of the antimony and arsenic still remaining, with a little 
copper and iron. Silica also dissolves. The following analy- 
ses are of air-dried Trail slime treated experimentally. In 
the first column is given the analysis of the air-dried slime, 



98 



LEAD REFINING BY ELECTROLYSIS. 



in the second column the same after treatment with ferric 

sulphate, and in the third column the same after treatment 
with HF solution: 

TABLE 39. 

H, O 14 . 5% 

Au 34 . 5 ozs. 36 . 44 ozs. 69 . 32 ozs. 

Ag. 15.9% 16.2% 31.9% 

Cu 9.5% .8% 1.28% 

Pb 16.0% 17.6% 33.1% 

Sb 25.91% 25.03% 3.72% 

As 5.96% 1.2% Nil. 

SiO, 2.2% 1.8% 0.8% 

The distribution of the products in this case were calcu- 
lated from the analyses to be about as follows: 

TABLE 40. 



In Copper-iron Solution. 


In Fluoride Solution. 


In Residue. 


No silver * 

No gold * 

No lead * 

92% of the copper 
8 . 5% of the antimony- 
Si % of the arsenic 
23% of the silica 


No silver * 

No gold * 

Xo lead * 
1% of the copper 
84.5% of the antimony 
19% of the arsenic 
59% of the silica 


100%, silver 

100%, gold 

100% lead 

7% of the copper 
6.75% of the antimony 
No arsenic 

18% of the silica 



* Known from tests of solutions. 



With the slime which has not been dried the results are 
somewhat better than the above, as the inevitable result of 
drying is the formation of hard lumps which it is hard for 
the solutions to penetrate. 

Usually silver dissolves, as a small excess of ferric sul- 
phate is always used, and before filtration the solution and 
suspended slime are agitated in presence of metallic copper 
until the excess of ferric sulphate has been reduced and all 
the silver removed from the solution. At a boiling temper- 



CHEMISTRY OF SLIME TREATMENT. 99 

ature tins takes from 2 to 10 hours according to the copper 
surface exposed, and the amount of ferric sulphate in excess. 
An exposure of about 4 square feet of copper for each cubic 
fool of solution is sufficient for moderately quick work. 

Settling and filtration is found to be much easier and 
quicker after silver has been removed. The solution settles 
clear in a short time and most of it can be drawn off without 
filtration. The residue may be washed by decantation to 
best advantage with hot water, or filtered in a press or on a 
horizontal cloth resting in a shallow tank with a perforated 
or grooved wood or lead bottom. Centrifugal machines are 
also used for this kind of work. If the material cools very 
much during filtration it clogs up from separation of anti- 
mony oxide. 

The extraction of silver requires a considerable excess of 
ferric sulphate, and even then with most slime the extrac- 
tion of silver is very incomplete. If the silver could be dis- 
solved and precipitated on copper in a separate tank, the 
expense of melting silver twice and parting would be saved. 
A temperature of 95-100° gives a much better extraction of 
silver than one of 80°. This was shown in • an experiment 
on slime containing 79% Ag; 12.6% Cu; 4.12% Sb; 88% Bi; 
3% Pb, from refining rich lead with 10% and 15% silver. 
The precipitated silver was washed in this case with HC1 to 
take out traces of Sb203 + Bi 2 03. A large excess of ferric 
iron was used, but it is doubtful if this made any great differ- 
ence in the result. The use of a considerable bulk of solu- 
tion had more influence probably. Particulars are given in 
Table 41. Lflfc 



100 LEAD REFINING BY ELECTROLYSIS. 

TABLE 41. 



No. 


Weight of 
Slime. 


Volume 
Solution. 


Fe" Used. 


! 1 

Time. Temperature. 1 Residue. 


1 
2 


45 gr. 
45 gr. 


2500 cc. 
2500 cc. 


50 gr. 
50 gr. 


3i H. 

3i H. 


80-85° 13 . 5 gr. 
95-99° 6.4 gr. 


No. 


Precipitated Silver. 


Contains Before Melting. 


Silver Extraction. 


1 
2 


25.7 gr. 
33.2 gr. 


98.71% Ag 
99 . 65% Ag 


71.4% 
93.0% 



The treatment of copper slime with ferric sulphate is very 
successful in removing copper quickly. With slime from 
blister copper anodes, there is too much sulphur present to 
allow of the solution of much if any silver. Several experi- 
ments have been made on slime analyzing Cu 53.29%; Ag 
12.90%; Bi 1.55%; Sb 3.30%: As 1.15%; S 11.96%; Te 
1.97%; Se .26%; Pb trace; gold and moisture not deter- 



mined. See Table 42. 


TABLE 


42. 






No. 


Slime Taken. 


Fe'". 


Temperature. Volume. 


H2SO4. 


Residue. 


1 
2 
3 

4 


200 gr. 161 gr. 
100 gr. 100 gr. 
700 gr. 750 gr. 
700 gr. 750 gr. 


90° 
85-92° 
85-90° 

85-92° 


1 200 cc. ! 
9 500 cc. ! 
10 000 cc. 


5%" 
4-2% 

4% 


122 gr. 

325 gr. 
1 See Table 
J 43. 



The residue was treated with caustic-soda solution to 
extract the sulphur, antimony, selenium, and tellurium if 
possible. Traces of tellurium dissolved out, but no selenium. 



CHEMISTRY OF SLIME TREATMENT. 
TABLE 43. 



101 



No. 


NaOH. 


Volume. 


Temper- 
ature. 


Residue. 


Fusion. 


Product. 


Agin 
Button. 


2 

3 

4 


30 gr. 
200 gr. 
200 gr. 


150 CC. 

2000 cc. 
2000 oc. 


Boiling 

1 1. 


19 . 7 gr. 
159.5 gr. 
177 gr. 


10 gr. nitre 
7.5 gr. soda 
90 gr. nitre 
100 gr. soda 
90 gr. nitre 
100 gr. soda 


Dore matte 
and slag 

Dore matte 
and slag 

Dore matte 
and slag 


12 . 22 gr. 
163 gr. 



Analysis of dore from 3 and 4, Ag 86.55%; Bi 5.37%; Cu 
5.99%; Au 1.62%; Te .16%; Se trace; Pb nil; Sb trace; 
Cu nil; As nil. 

The residue from No. 1 was melted direct to matte, with- 
out treatment with NaOH. Matte weighed 48 grams. Con- 
tained 12.7% S; 53.6% Ag calculated. Probably a great 
deal more caustic soda was used than was entirely necessary. 
Probably 80 grams of caustic for 700 parts slime taken would 
have done just as well. Milk of lime would act similarly to 
caustic soda and be cheaper. 

In treating copper slime with ferric sulphate, the process 
works quickly and completely at a temperature of about 
90°; at 100° the liberated sulphur sticks together and hinders 
the reaction. Only a slight excess of ferric iron should be 
used, and the excess reduced by suspending copper plates 
in the solution before removing it from the insoluble residue. 

Returning to the consideration of lead slime-treatment, 
the solution, after removal from the slime, now containing 
ferrous sulphate, cupric sulphate, arsenious acid, and sul- 
phuric acid, beside smaller quantities of arsenic, bismuth, 
silica, and fluosilicic acid, is to be electrolyzed for metallic 
copper and regeneration of ferric sulphate. A separate treat- 
ment of the solution with copper oxide, metallic copper and 



102 LEAD REFINING BY ELECTROLYSIS. 

air, or copper matte, is necessary, unless the slime being 
treated should have been air-dried, and say two-thirds oxi- 
dized. 

The electrolysis of the solution takes place at about 40°, 
and on cooling to this temperature or a little lower about 
10 grams antimony oxide per litre and excess of arsenious 
acid above that required to saturate the solution at this tem- 
perature (say 2% As 2 3 ) crystallize out. This cooling takes 
care of the arsenic of the slime, the solution, after reaching 
a concentration of about 2% As 2 3 , thereafter depositing that 
removed from the slime. The arsenic crystallizes as bright, 
hard crystals. The solubility of As 2 3 in the hot solution 
is about one part in ten parts solution, and at 20° one part 
in 100 parts solution, having therefore a large variation for 
difference in temperature. There are two varieties of AS2O3, 
but we have here to do, at least in the cold, with the crystal- 
line variety, of which the solubility is ten parts in 100 parts 
hot water and 1.7 parts in 100 parts cold water (Comey's 
Dictionary). The solubility in reduced iron solution is not 
very different. 

The electrolysis of solution containing iron and copper 
for the production of a copper deposit and a solution of ferric 
sulphate was first proposed by Body.* 

Siemens and Halskef proposed a process in which the fer- 
ric sulphate was used to attack metallic copper and copper 
sulphide and the solution then brought back to the electrolytic 
cell for the recovery of the copper and the ferric sulphate. 
Difficulties were met by Siemens and Halske in the electrolysis, 

* U. S. A. Patent 338150. Jan. 5, 1886. 

t German Patent 42243. Sept. 14, 1886. English Patent 14033. Nov.. 
1, 1886. 



CHEMISTRY OF SLIME TREATMENT. 



103 



particularly the carbon anodes wen 4 corroded and the yield 
of ferric sulphate was low. The corrosion of the carbon 
anodes was a fatal difficulty. I found that the anodes could 
be made to last permanently if they were kept in constant 
motion through the solution.* 

The electrolysis of the reduced iron solution has been 
made the subject of a special study to determine the effect 
on the current density and voltage of variations in temper- 
ature and chemical composition. 

The electrodes used in the test were each of graphite, and 
the anode was kept in back-and- forth motion through the 
electrolyte by means of a crank. If the anode stopped it 




u 21 

Amperes per Square Foot 
Fig 11. — Solution 0. 



polarized in a short time, and oxygen was evolved on the 
anode and little or no ferric iron formed. As the anode reac- 
tion was the only one with which difficulty w^as experienced 
before the requirements of the case were understood, the depo- 
sition of copper at the cathode w T as disregarded, and a solu- 
tion electrolyzed containing ferrous sulphate, copper sul- 



* U. S. A. Patent 803543. Nov. 



1905. 



104 



LEAD REFINING BY ELECTROLYSIS. 



phate, and sulphuric acid, and in some cases also ferric sul- 
phate, without a diaphragm. 

The results indicate that the effect of temperature is the 
most important. The results are plotted as Figs. 11 to 17. 
The ordinates represent the polarization in excess of the 
electromotive force required to carry out the oxidation of 
the iron. 

Solutions were tested as follows: 

TABLE 44. 



Solution. 


O 


A 


B 


C 


D 


HaSO* grams per 100 cc 

FeS0 4 7H 2 " " 100 " 

CuS0 4 5H 2 " " 100 " 


1 

5 

12 


2 

5 

12 


3 

5 

12 


5 
5 

12 


9 

5 

12 







The amperes per square foot refers to plane occupied 
by 1 inch carbon rods spaced l^J inch centre to centre. 
For amperes per square foot of carbon surface, multiply 
by 1.09. 

Tests were also made with the following solution: 

TABLE 45. 



Solution. 



H 2 S0 4 grams per 100 cc 

FeS0 4 7H 2 " " 100" 
CuS0 4 5H 2 " " 100" 
Fe 2 (S0 4 ) 3 " " 100" 



O' 


A' 


B' 


C 


1 


2 


3 


5 


5 


5 


5 


5 


4 


4 


4 


4 


10.7 


10.7 


10.7 


10.7 



D' 



9 

5 

4 

10.7 



The results are somewhat different, probably on account 
of the failure of copper to deposit on the cathodes in the 
second series where the reduction of ferric iron takes place 



CHEMISTRY OF SLIME TREATMENT. 



105 




7 14 21 

Amperes per Square Foot 

Fig. 12. — Solution A. 



1.0 













Is 
1! 




•+■ 










b d 










3~ 










o 










Pi 











14 21 

Amperes per Square Foot 

Fig. 13. — Solution B. 




14 21. 

Amperes per. Square Foot 

Fig. 14. — Solution C. 



106 



LEAD REFINING BY ELECTROLYSIS. 



1.0 




* _ 

* 


J 


■= -i-^-25- 










i *- *55° 




.5 


Is 










■~a 
























o 












fk 











14 21 

Amperes per Square Foot 

Fig. 15. — Solution D. 



1.0 






^— ^r 


10.5 amps. 
L ps. 


jfc 1 








'/• i4 am 




la ) 












.5 


C-i-i 


> — ="- — — 


n^ 


$f 







1% 2X 3% 5% 0% H 2 So 4 

Fig. 16. — Effect of Sulphuric Acid at 25° C. 




n 



2% Z% 



5% 
Per Cent H2SO4 

Fig. 17.— Effect of Sulphuric Acid at 50°-55° C. 



CHEMISTRY OF SLIME TREATMENT. 



107 



instead. I regard these latter results as showing the anode 
polarization best. See Figs. IS, 19, 20, 21. 

The necessity of moving the anodes exists under the con- 
ditions studied in these experiments if polarization is to be 
prevented. However, I found that at a still higher tem- 
perature, near boiling, it is no longer necessary to move the 
anodes.* 

Of course, at lower temperatures, the anode rods might 
be left stationary if the relative motion between anode sur- 
face and electrolyte was maintained by rapid circulation, 



1.0 





#10 










3 -. 
.2-2 




20° 


x 






o 

to 






( 


^0° 





14 21 

Amperes per Square Foot 

Fig. 18. — Solution A'. 



.28 



but it would have to be so rapid as to be impracticable on a 
large scale, in a tank of any ordinary construction. 

It was found in some experiments on slime treatment 
that the anodes polarized in spite of everything that could 
be done, including increasing temperature and the velocity 
of the anodes. The anodes on taking out were slimy to the 
touch; after brushing off they would run some hours suc- 
cessfully and would then polarize again. 



* U. S. Patent applied for. 



108 



LEAD REFINING BY ELECTROLYSIS. 



1.0 



.5-S 



is 




f 


/45° 


S.S 

o 

Pi 






ORO y 








• 







14 21 

Amperes per Square Foot 

Fig. 19.— Solution B'. 



1.0 



















*22° 




o 

*38 


o 

50- 


CB'lH 




/ 












ij 


L— — «— — — "* 







14 21 

Amperes per Square Foot 

Fig. 20.— Solution C 



1.0 



la 




20°/ 


0/ 

37/ 


mV 


o 



















14 21 

Amperes per Square Foot 
Fig. 21.— Solution D\ 



CHEMISTRY OF SLIME TREATMENT. 100 

As the process has been worked continuously on other fer- 
rous sulphate solutions than those from treating lead slime, an 
investigation was made to ascertain the cause of the trouble. 

Pure solution of iron and copper sulphates and sulphuric 
acid were treated with various materials and electrolyzed. 
The presence of gelatine, tin, arsenic, antimony, bismuth, 
and soluble silica had no prejudicial effect. 

Fluosilicic acid, on the other hand, caused polarization 
readily, and if the quantity added was at all large, a thick 
silica deposit would form on the anode. The coating from 
anodes used in working up solution from slime treating was 
tested and found to consist largely of silica. 

For large scale w T ork, the remaining serious question is 
oik 4 of diaphragms. For this process diaphragms of wood, 
about |" thick with §" to J" holes bored through as closely 
as possible, with holes filled with wet asbestos; asbestos 
boards \" thick, hardened by absorption of the right amount 
of sulphur; and pairs of perforated lead sheets with several 
thicknesses of asbestos between have been tried, and all 
have given success. 

The disadvantage of the wood diaphragms has been that 
the plugs, if not put in tightly enough, drop out, or if the 
copper deposit gets spongy, which has happened when unre- 
duced solution was fed in, the copper may grow into the plugs 
and on drawing the cathodes, a plug or plugs come too. 

The disadvantage of the asbestos board hardened with 
sulphur is that it expands slightly when wet and warps. 
This difficulty, I believe, can be cured by soaking the boards 
a week or two before putting them in a tank. The resistance 
is quite a little higher, requiring perhaps .4 to .5 volt more 
to operate a tank than one with lead and asbestos diaphragms. 



110 LEAD REFINING BY ELECTYOLYSIS. 

The disadvantage of the lead and asbestos boards dia- 
phragm is the cost of the lead, and the necessity of operating 
the tank at a uniform temperature to prevent wrinkling of 
the lead. 

These advantages would have to be weighed against each 
other before making a choice, but good success will result in 
the use of any. 

To prepare hardened asbestos diaphragms of the above 
construction, asbestos " mill boards," which come in about 
40-inch squares, should be placed flat on a floor, powdered 
sulphur sprinkled on evenly, and placed in an oven hot 
enough to melt sulphur, for an hour or more. The sulphur 
melts and is absorbed by the asbestos. The same operation 
is repeated on the other side of the boards. The hot board 
is cooled on a flat floor, giving a sheet of considerable stiff- 
ness and strength, that does not soften in water or acid solu- 
tion, even after a long time. Care must be used not to fully 
laturate the board with asbestos, which would make it an 
insulator. The effect of the sulphur is to cement the asbestos 
fibres together. Two to three pounds of sulphur is found to 
be about right for 10 square feet of \" board. 

For details of construction of lead-lined copper-iron sul- 
phate electrolytic tanks, see Chapter VII. The catholyte 
only comes in contact with the lead lining in these construc- 
tions. The solution of ferrous and cupric sulphates and sul- 
phuric acid, containing approximately 30 grams copper, 40-50 
grams ferrous iron and 20-60 grams H 2 S0 4 per litre, is fed 
in a continuous stream into the cathode compartment, which 
stands in composition at about 10 grams copper, 40-50 grams 
ferrous iron and 20-60 grams H 2 S0 4 . An overflow about two 
inches below the top of the tank is provided for the anolyte, 



CHEMISTRY OF SLIME TREATMENT. HI 

averaging S-lO grams ferrous iron, 30-40 grams ferric iron, 
20-60 grams IL.SO., per litre. The effect of feeding solution to 
the catholyte is to maintain the catholyte at a slightly higher 
Level than the anolyte, so that the solution percolates through 
the diaphragm continuously, preventing back-flow of anolyte 
to the catholyte compartments. The anolyte is also found 
to be slightly heavier than the catholyte, for instance, 1.19 
and 1.16 specific gravity respectively. 

The tanks are built on the principle of placing a series 
of anolyte boxes, with catholyte spaces between each, and 
on the sides and bottom too. Circulation of the catholyte 
through the tank can be easily arranged and circulation of 
the anolyte is provided by siphons connecting each anolyte 
compartment to a trough on each side of the tank. This 
trough need not necessarily be placed outside of the tank, 
but can be fitted inside. Circulation is maintained by com- 
pressed air. The trough on one side serves as a feed to all 
the anolyte compartments, and the discharge takes place 
to the trough on the other side. The siphons are provided 
with an arrangement by which the air can be sucked out. 

Serious attempts were made to electrolyze the solution 
in cells without a diaphragm, depending on the formation 
of a heavier ferric sulphate solution at the anode, which 
should settle to the bottom of the cell. This principle can 
be applied successfully in electrolyzing chloride solutions, 
but it will be difficult or impossible, I think, to use it in 
iron sulphate electrolysis. 

The slime after treatment with ferric sulphate should be 
washed fairly well, as any iron and copper salts not washed 
out will accumulate in the fluoride solution used for anti- 
mony extraction. Copper can, however, be removed by 



112 



LEAD REFINING BY ELECTROLYSIS. 



antimony as described in Chapter III, and the only effect 
of iron is to slightly diminish the current efficiency of the 
antimony deposition, but not very seriously. 

The fluoride solution dissolves most of the antimony pres- 
ent, as well as the arsenic still remaining, and traces of bis- 
muth and silica. 

The solubility of bismuth in the fluoride solution, provided 
excess of acid is used, is very slight. Considerable quantities 
dissolve, if no excess of HF is used, and HF added to the solu- 
tion in that case causes a precipitation of bismuth, probably 
as BiF 3 . The amount of bismuth dissolved with excess of 
HF present has been variously determined from .008 to .010 
grams per 100 cc. 

The bismuth dissolved deposits out with the antimony, 
and on one occasion, treating high bismuth slime, the per- 
centage of bismuth in the antimony was 0.67. This is the 
highest percentage yet observed, and is equivalent to .035 
grams Bi dissolved per 100 cc. 

The extraction of antimony with HF from the slime after 
treatment with ferric sulphate averages 95%, a tempera- 
ture of 30-40° C. and excess of HF being desirable. The 
effect of H 2 S0 4 , which is also present in the solutions, seems 
to be insignificant. See Table 46. 

TABLE 46. 



Weight of 
Slime. 


HF 

Excess. 


Weight of 
Residue. 


Tempera 
ture. 


| 
Slime. Residue. 


Extracted, 
Sb. 


200 gr. 
100 gr. 

50 gr. 

50 gr. 


17% 
200% 
200% 
200% 


107 
50.2 
26. 
' 26.2 


20° C. 

20° C. 
35-40° C. 
30-40° C. 


30 . 8% Sb! 5 . 59% Sb 90 . 3% 
30 . 8% Sb! 3 . 99% Sb j 93 . 5% 
30.8% Sb! 2.38% Sb ! 96.1% 
30 . 8% Sb 3 . 63% Sb 93 . 8% 



CHEMISTRY OF SLIME TREATMENt. 113 

No silver dissolve's in the fluoride solution, probably on 
account of the presence of other unoxidized metals capable 
of precipitating silver. The antimony fluoride solution is 
treated with K L »S0 4 or Na 2 S0 4 for removal of SiF 6 and elec- 
trolyzed for metallic antimony and regeneration of HF. Par- 
ticulars will be found on page 144. 

The treatment of the insoluble residue has only been carried 
out by fusion to dore bullion. This fusion can be accom- 
plished with various fluxes, but soda has been chiefly used 
as a flux in the experiments, which was a mistake. Fusion 
with silica is better, and gives a clean dore bullion. The 
sulphur and carbon in the slime are oxidized by the oxygen 
liberated when lead sulphate is decomposed by silica. The 
lead silicate slag can be smelted for the lead and traces of 
silver it contains. 

The fusion may be conducted in reverberatories, or cru- 
cibles, though the latter is best, for no furnace refining is 
required. 

A sample of dore bullion produced from Trail slime, con- 
taining in the first place approximately 30% Sb; 29% Ag; 
6% As; 10% Pb; and 7% Cu, which had been treated with 
ferric sulphate and hydrofluoric acid, and then melted with 
soda, contained Ag 78.94%; Pb 17.56%; Au 2.08%; Cu .81%; 
Sb .47%; no As. Other melts with silica have produced far 
cleaner dore, containing, beside gold and silver, only traces 
of copper and lead. 

The metallurgical recovery of the ferric sulphate process 
is excellent. 95 lbs. of Trail slime contained by corrected 
fire assay 445.83 ozs. silver and 3.7 ozs. gold. This was 
treated experimentally in some 8 or 10 batches, using the 
solutions over and over again, and notwithstanding some 



114 LEAD REFINING BY ELECTROLYSIS. 

accidents, the silver recovery was 443.85 ozs. and gold 3.65 ozs. 
The limit of accuracy of the gold assays was .1 oz., so prob- 
ably the actual recovery of gold was as great in proportion, 
or greater than that of silver. The silver loss was less than 
§%, and on the basis of uncorrected assay there would have 
been a gain of from |% to 1%. 

Copper scale for adding copper and oxygen to the iron 
sulphate solution is not to be recommended, as it contains 
too much metallic copper and cuprous oxide. Copper sul- 
phate is rather too expensive, though it may only represent 
roasted copper matte plus sulphuric acid. The use of granu- 
lated copper in a tower, through which the acid slowly passes 
in the presence of air is permissible but slow, requiring a large 
stock of metallic copper. The copper is relatively more 
expensive than the same material in the form of roasted 
matte. 

Methods of treating roasted copper matte for extraction 
of copper are well-known, the best description being that 
given by Hofmann.* The material treated at Argentine, 
Kansas, contained 40% Cu and 12-14% Pb. It was ground 
to 50 mesh in a ball-mill, and roasted in Pearce turret fur- 
naces. The roasted material was again ground to 50-mesh 
in a ball-mill and treated in tanks with stirring-gear, with 
water and sulphuric acid. The mixture was filtered in a 
wood filter-press and the solution treated with a further small 
quantity of matte while air was blown through to purify 
the solution from iron, arsenic, and antimony. 

The air blowing can be omitted in slime treating, as the 
presence of ferrous iron in the solution is not an objection. 

* Mineral Industry, Vol. 10, page 231. 



CHEMISTRY OF SLIME TREATMENT. 115 

One treatment of the solution with a slight excess of matte 
would be sufficient. The arsenic, antimony, and fluosilicic 
acid being removed by neutralization, the result is a neutral 
solution of cupric and ferrous sulphates. This requires 
acidification to say 2% H 2 S0 4 , before electrolysis, to save 
power. 

The insoluble residue would contain considerable anti- 
mony, and if the solution contained traces of bismuth, con- 
siderable of that beside a good deal of lead and some copper. 
By smelting the leached matte in a lead-furnace the anti- 
mony and bismuth values would be recovered in the lead pro- 
duced. On the supposition that 59 lbs. ferric iron are re- 
quired to treat the slime from one ton of lead, which is a fair 
average, and that the matte contains 40% copper and 14% 
lead, the lead bullion produced by smelting the leached matte 
alone would contain as much as 20% antimony, and bismuth 
up to 16%, if bismuth is present in the slime in large quan- 
tity. This bullion could be refined for the lead content 
without difficulty in the usual way, and the slime treated 
with dilute nitric acid to make bismuth subnitrate and anti- 
mony oxide. It would probably be more advantageous to 
dilute the matte with lead ore before smelting to produce a 
purer bullion with less loss of antimony and bismuth in the 
furnace. 

If the lead bullion only contained a little bismuth, as is 
usually the case, say i to 1 lb. per ton lead, the bismuth 
would be practically all recovered from the leached matte, 
in the resulting lead bullion. 

Perfluoride processes. — Antimony pentafluoride, and also 
ferric sulphate with the addition of HF amounting to the 
use of ferric fluoride, have been tried. The ferric sulphate 



116 LEAD REFINING BY ELECTROLYSIS. 

and HF process possesses the advantage over the ferric sul- 
phate process, that the antimony goes into solution with the 
copper and arsenic. At the time the experiments were made, 
I was trying to dissolve silver with the copper, arsenic, etc. 
Silver is not dissolved nearly as well in presence of HF by 
ferric sulphate as without HF. The experiments were given 
up on account of the inability of dissolving silver, but if this 
was not required, and it was possible to separate the arsenic 
by crystallization as AS2O3, and a diaphragm cell, unattack- 
able by HF, could be provided for electrolyzing the solutions, 
the process would be workable as well as simple and quick. 
The operations ought to be treatment of slime with solution 
of copper, antimony, and arsenic. Insoluble residue con- 
sists of lead sulphate, silver, gold, and bismuth fluoride. In 
solution, ferrous sulphate, cupric sulphate, antimony trifluoride 
and arsenious acid, .01-02% bismuth, and stannic fluoride, 
in case slime contains tin. 

The solution would then be electrolyzed with antimony 
anodes and copper cathodes, with a current density of 3-5 
amperes per square foot until nearly all copper was deposited 
out. Then a short electrolysis with antimony anodes and 
copper cathodes in a separate cell would remove the remain- 
ing copper with some antimony. The solution would then 
be electrolyzed for the metallic antimony and the regeneration 
of the ferric salt, and cooled at some stage of the process to 
crystallize out AS2O3 if possible. 

The slime treated contained Ag 29.2%; Cu 7.1%; Pb 
10.2%; Sb 30.5%; As 6.10%; O 6%; H 2 not deter- 
mined. 

Some of the results are given in Table 47. 



CHEMISTRY OF SLIME TREATMENT. 
TABLE 17. 



117 



Slime. 


HF. 


H-jSOi. 


Fe'". 


Tempera- 
ture. 


Volume. 


Dissolved. 


Fe"' 

Excess. 


100 gr. 
50 gr. 


30 gr. 
15 gr. 


200 gr. 
85 gr. 


130 gr. 
30 gr. 


Hot 


4000cc. 
700 cc. 


10.5 

None 


75 



The filtrate from the second treatment was boiled up with 
fresh slime to throw out copper and arsenic, and electrolyzed 
with lead cathode, CD. 9-18 amps, per square foot, and lead 
anode CD. 54-108 amps, per square foot. The antimony 
deposited contained 1.62% Cu and 5.85% Pb (from cathode). 

The process was varied by treating unoxidizeel slime, with 
ferric fluoride and sulphuric acid, in quantity sufficient to 
extract antimony, and then with ferric sulphate alone to 
extract copper. 

The slime had the same analysis as the above, but was 
reduced by treatment with lead and fluosilicic acid to get it 
back to its original metallic condition as nearly as possible. 
The first solution contained in 3240 cc. 76 grams Fe'", 
125 gr. HF, 100 gr. H 2 S0 4 . Solution after the reaction 
contained Cu 4.68 grs.; Sb 47.35 grs.; As 8.02 grs. On 
standing in a lead pan all the copper deposited out, as well 
as a small amount of antimony on the lead. 

The second solution applied to the slime contained in 
2800 cc. 110 grs. Fe ,// and 150 gr. H 2 S0 4 . After reaction, 
the solution contained 47^25 grams of silver, precipitated 
out by metallic copper, while 41 grams copper dissolved. The 
solution then contained Ag, none; Cu, 50.9 gr.; Sb, 2.14 gr.; 
As, 3.12 gr. The residue contained PbS0 4 , 63.5%; Pb, 2.42%; 
Cu, none; Sb, 6.24%; As, .5%. 

The results are given in Table 48. 



118 



LEAD REFINING BY ELECTROLYSIS. 
TABLE 48. 



In Slime. 



In Fluoride 
Solution. 



In Sulphate 
Solution. 



In Residue. 



Silver. . . . 
Copper. . . 
Arsenic. . . 
Antimony- 
Lead 



58 . 4 grs. 
14.2 grs. 
12.2 grs. 
61.0 grs. 
20.4 grs. 



4.68 

8.02 

47.35 



47.25 gr. 
10.30 gr. 

3.25 gr. 

2.23 gr. 



None 
.49 gr. 
6.17 gr. 
45 gr. 



The electrolysis was intended to be carried out as fol- 
lows: The fluoride solution was to be electrolyzed for ferric 
fluoride and antimony, and the sulphate solution for ferric 
sulphate and copper. 

Antimony pentafluoride. — This process is intended to dis- 
solve everything from the slime except gold, lead, and bis- 
muth, the last two of which are insoluble. Antimony penta- 
fluoride was thought to be a stronger oxidizer than ferric 
fluoride. 

I electrolyzed antimony trifluoride solution containing 
about 14% Sb as SbF 3 , freed from H 2 SiF 6 by adding KF, to 
precipitate K 2 SiF 6 , with a graphite anode and lead cathode, 
separated by cotton cloth. The e.m.f. required to carry out 
the reaction, 

5SbF 3 = 3SbF 3 + 2Sb. 



is about 1.45. The polarization was about .2 volt. The 
current density in my experiment varied from 21 to 40 am- 
peres per square foot on cathodes, and a little less on anodes, 
with an e.m.f. with the latter current density of 2.35 volts. 
The process is not a success because frequently it is difficult 
to reduce the SbF 5 formed, and its action on slime is far too 
slow. 



CHEMISTRY OF SLIME TREATMENT. 119 

Ferric salts of strong monobasic acids as oxidizers. — Fer- 
ric acetate was tried and found to be valueless. On the 
other hand, with strong acids (see page 19), especially methyl 
sulphuric acid (for the preparation of which sec Chapter IV), 
dissolves from the slime at one treatment, bismuth, copper, 
arsenic, and lead, leaving silver, gold, and antimony trioxide. 
HF precipitates insoluble bismuth fluoride from the solution, 
copper is precipitable by lead, and the solution of ferrous 
and lead methyl sulphates may be electrolyzied for ferric salt 
and lead. In distinction to ferrous sulphate (polybasic acid), 
ferrous methyl sulphate is easily oxidized with a carbon 
anode at the ordinary temperature, even though the anode 
is not moving. The difference is probably clue to difference 
of valency. 

The reaction Fe (S0 4 CH3)2 + S04CH3 , = Fe(S04CH3) 3 is a 
simpler reaction than 

2FeS0 4 + 2HS0' 4 = Fe 2 (S0 4 ) 3 + H 2 0. 

In the first case the anion reacts with a molecule present in 
large quantity, while in the second case the reaction requires 
the molecular connection or contact of four different parts, 
which can readily be conceived to occur less often. 

Use of lead peroxide as oxidizing agent for slime. — If lead 
fluosilicate solution is electrolyzed with a carbon anode and 
a lead cathode a solid smooth coating of Pb0 2 is deposited 
on the anode, and if the solution contains gelatine a smooth 
deposit of lead is deposited on the cathode. The lead per- 
oxide in its massive form is quite inactive, but if ground fine 
and mixed with raw slime in the presence of fluosilicic acid 
the metals of the slime will be oxidized into solution. 



120 LEAD REFINING BY ELECTROLYSIS. 

Lead peroxide was mixed with slime and lead fluosilicate 
flue-silicic acid electrolyte for the purpose of extracting lead, 
copper and silver and leaving a residue of antimony trioxide 
and gold. The results were not satisfactory either in point 
of time required or in extraction of metals. 

Later experiments showed the possibility of having suffi- 
cient HF present to take all the antimony into solution 
along with the other metals. A solution of lead fluosilicate 
containing, for example, 5-6% lead and 15% SiF 6 , permits 
of* the addition to it in the cold of about 5% anhydrous HF 
without causing a precipitation of lead, and at a higher tem- 
perature considerably more ma} r be added. The explana- 
tion of this is that fluosilicic acid is a considerably stronger 
acid than hydrofluoric acid and is capable of decomposing 
insoluble lead fluoride until the percentage of hydrofluoric 
acid becomes great enough to precipitate lead fluoride and 
a condition of equilibrium is reached. On the other hand 
a solution of antimony trifluoride may be added in any quan- 
tity to the lead fluosilicate solution, without causing precipi- 
tation of lead fluoride, consequently it is feasible to take the 
antimony into solution simultaneously with lead, by having 
a certain amount of hyrdofluoric acid present. Furthermore, 
the recovery of antimony from a mixed solution of antimony 
fluoride and lead fluosilicate can be nicely carried out by 
electrolysis with a lead anode and a carbon cathode. Anti- 
mony deposits on the cathode and lead fluosilicate dissolves 
on the anode until the percentage of hydrofluoric acid in the 
solution becomes quite high, and thereafter lead precipitates 
as PF 2 in the neighborhood of the anode. On these prin- 
ciples I thought a good slime process could be based, but the 
experiments have not been entirely successful so far, presum- 



rilKMlsTRY OF SLIME TREATMENT. 121 

ably on account of the formation of antimony pentafluoride, 
from the reaction of lead peroxide and antimony fluoride. 
At any rate the antimony goes into solution from the slime 
in an irreducible form. 

Thirty-three gr. air-oxidized Trail slime containing about 
15.8% Ag; 8.2% Cu; 16.0% Pb; 26.0% Sb; 5.96% As, was 
treated with 150 cc. H 2 SiF 6 and 17 cc. of 50% HF, and 50 cc. 
of water, and 25 gr. finely-ground electrolytic Pb0 2 added. 
The solution warmed up quite a little when Pb0 2 was. added, 
showing a rapid reaction. About half the silver went into 
solution with practically all of the other metals, except some 
arsenic. Had the slime not been air-oxidized much more 
Pb0 2 would have been required. 

Silver was removed by precipitation with copper, and 
the solution electrolyzed with a carbon anode and copper 
cathode, for recovery of Pb0 2 , and metal. Until most of 
the copper had been removed, a good copper deposit was 
obtained. Then the cathode darkened and eventually the 
deposit evidently consisted of lead. It contained no anti- 
mony, showing the presence of an antimony compound widely 
differing from the ordinary variety. 

In another experiment unoxidized specially prepared 
slime, containing on dry sample Ag 4.5%; Bi 1.1%; Cu 17.4%; 
Sb 38.0%; As 12.0%; Pb 11.0%, was treated with lead 
electrolyte containing about 4% Pb and 20-25% SiF. The 
solution had been prepared by the electrolysis of a solution 
high in lead and containing some HF, though not enough 
to precipitate lead at any time. This amount of HF was 
sufficient for the experiment, so none was added. On adding 
the finely-ground Pb0 2 necessary for the reactions given 
below the temperature rose rapidly. 



122 LEAD REFINING BY ELECTROLYSIS. 

Cu+ Pb0 2 +2HSiF 6 =CuSiF 6 +PbSiF a +2H 2 0; 

Pb+ Pb0 2 +2H,SiF 6 =2PbSiF 6 + 2H 2 0; 

2Sb + 3Pb0 2 + 6HF + 3H 2 SiF 6 = 2SbF 3 + 3PbSiF 6 + 6H 2 ; 

2As + 3Pb0 2 + 3H 2 SiF 6 = AS 2 3 + 3PbSiF 6 + 3H 2 ; 

2Ag+ Pb0 2 +2H 2 SiF 6 =Ag 2 SiF 6 +PbSiF 6 + 2H 2 0; 

2Bi + 3Pb0 2 + 6HF+ 3H 2 SiF 6 = 2BiF 3 + 3PbSiF 6 + 3H 2 0. 

The actual increase of temperature was 15°, while the 
energy of the reaction was about equivalent to a change of 
temperature, allowing something for the box, of about 26°, 
so the reactions actually taking place only amounted to 
57.5% of the total energy expected. At the time this was 
thought to be on account of not entirely completed reac- 
tion, which is no doubt the case to a considerable extent, but 
the formation of an antimony or arsenic compound of higher 
valence is also probable. The residue consisted of 30% by 
weight of the orginal dry weight. 

By heating the residue to a high temperature with the 
solution further reactions took place, with the solution of 
some of the slime, and reduction of the solution. 

The advantages of the process would be important, men- 
tioning : 

(1) The slime need not be washed or even removed from 
the electrolytic tanks, as the slime solution and lead refining 
solutions are the same, and the excess accumulating in the 
slime plant, derived from the electrolyte contained in the 
slime treated, would be returned, after proper purification. 

(2) The metals are directly recovered by electrolysis in 
a good state of purity. 

(3) The electrolytic tanks are of the simplest kind, no 
diaphragms being necessary. 

The chief disadvantages would be the necessity of col- 
lecting a gpod deal of Pb0 2 and grinding it, and the necessity 



CHEMISTRY OF SLIME TREATMENT. 123 

shown to exist of working the slime treatment at a high tem- 
perature. 

The electrolytic deposits obtained would consist first of 
copper and then of an alloy of copper and antimony, then of 
antimony, then of impure lead containing mostly arsenic and 
some antimony. The intermediate products may be refined 
in the same solution, using the impure cathode as anode in 
a separate cell through which the solution passes at the 
appropriate stage of its progress through the tanks. For 
instance, the antimony copper alloy deposited intermediately 
between pure copper and rather pure antimony, would be 
refined in the solution which contains copper as it first comes 
to the electrolytic tanks. In this way the impure cathodes 
would not accumulate, but a certain quantity would always 
be on hand in the course of working up into pure metal. In 
the example given the copper and antimony of the alloy 
dissolve at the anode, while only copper deposits at the 
cathode and the antimony accumulates in the solution. The 
power consumption per ton of bullion of an ordinary quality, 
containing say 1% antimony, \% each copper and silver, 
and t 3 o% arsenic would amount to about 45 K.W. hours, 
which is very moderate. In fact the power requirement is 
less than in any other electrolytic slime process discussed 
so far. 

Alkaline regeneration processes. — Alkaline solutions con- 
taining sulphides are the only ones that will dissolve much 
from the slime. Hypochlorite solutions were tried, and 
arsenic was removed quite well, but it had not much action 
on anything else. Unsuccessful preliminary trials were also 
made with hyposulphite solutions also containing tetrathion- 
ate. 



124 LEAD REFINING BY ELECTROLYSIS. 

Slime suspended in sodium sulphide and air drawn 
through gives up the antimony and arsenic readily.* Air 
oxidation is much more efficient with an alkaline solution 
or a solution of a monobasic acid as HC1, than with the 
customary sulphuric acid. Since the heat of combination of 
sulphur (liberated by oxygen and dissolved in the solution) 
with antimony and arsenic to form sulphosalts is probably 
greater than that with copper and silver, it would be expected 
that a good extraction of antimony and arsenic could be 
obtained without forming much silver and copper sulphides, 
though the lead would probably be converted to sulphide. 

The electrolysis of sulphantimonite solutions is described 
by Borchers.| The yield of antimony is quantitative on 
amount present but not on current used. The anode reac- 
tions were the liberation of sulphur which combined to form 
poly sulphides, and the formation of sodium hyposulphite. 
The polysulphide would be of immediate use as solvent for 
antimony in a following slime treatment, but the formation 
of Na 2 S 2 3 represents at least a temporary loss. The per- 
centage of the current employed in the most desirable reac- 
tion for our purpose, namely, 

(1) Sb 2 S3 + 3Na 2 S = 2Sb + 3Na2S2 is calculated from Bor- 
cher's figures to be 35.8% in both cases given, and that 
employed in the reaction. 

(2) 4Sb 2 S 3 + 9H 2 + 12Na 2 S = 8Sb + 3Na 2 S 2 3 + 18NaSH fig- 
ures 80% or over in the first case, and in the second case, 
about 80%. This shows that some hydrogen was liberated 
on the cathodes in place of antimony, as in fact must have 
been the case, to get all the antimony out. 

* Results at Trail show arsenic to be mostly insoluble. 

f Electric Smelting and Refining. Second Eng. Ed., page 476. 



CHEMISTRY OF SLIME TREATMENT. 125 

The relative proportion of the most desirable reaction 
(1) and the undesirable reaction (2) is shown to be about 
3195 and 69% of the total. 

What to do with the arsenic accumulating in the solution 
is another question to be considered. The current efficiency 
in depositing antimony is evidently rather low, unless 
diaphragms are used * A diaphragm of asbestos, supported 
between perforated iron plates, would be analagous to the 
same 1 construction using lead instead of iron, which is entirely 
satisfactory in ferric sulphate electrolysis. 

The conversion of thiosulphate back to sulphide could 
be effected by evaporating to dryness and igniting with car- 
bon, removing oxygen and water from the mixture of NaSH, 
Na 2 S 2 , NaOH, and Na 2 S 2 3 . 

That no great difficulty would be met in treating the 
insoluble portion of the slime, even if converted to sulphide, 
by fusing to matte and heating the ground matte with concen- 
trated sulphuric acid, is evident from the description given 
on page 78. 

Treatment with copper fluosilicate. — As copper stands 
below arsenic, antimony, bismuth, and lead in the e.m.f. 
series for fluosilicate solutions, it was thought that treat- 
ment of slime with copper fluosilicate solution containing 
some HF would result in the solution of the above metals 
with a precipitation of the corresponding amount of copper, 
while the residue would be treated for copper and silver by re- 
fining, and the solution for arsenic, antimony, lead and bis- 
muth in the same manner as described on page 135. No 



About 40% efficiency at Trail. 



126 LEAD REFINING BY ELECTROLYSIS. 

reaction takes place, however. The addition of HF does 
not help the result. 

Compression of slime to an anode plate for direct elec- 
trolysis. — In many ways this seems the most logical method 
of all. We then have merely a complicated electrolytic refin- 
ing operation to conduct. This it is, however, possible to do. 
The appropriate solution to begin with, is a solution of cop- 
per fluosilicate, fluosilicic and a few percent of hydrofluoric 
acid. At the anode lead, arsenic, antimony, bismuth, and 
copper dissolve, while copper deposits on the cathode. 
Fresh copper solution is continuously required, while a solu- 
tion containing lead, arsenic, antimony, and a little bismuth 
and copper is produced. This can be worked up to the stage 
of containing only a little antimony and arsenic beside very 
much lead, in the same manner as described on pages 136 
and 137. 

The remaining step is the electrolysis of the solution with 
a lead cathode and copper anode in an electrolytic cell, with 
a diaphragm for the production of lead on the cathode and 
copper fluosilicate solution at the anode. This can also be 
done in a gravity cell with a horizontal lead cathode above 
and copper anode underneath. 

Oxidizing slime suspended in solution by air-blast. — At 
Trail the first method of slime treatment consisted in blow- 
ing air through the slime suspended in H 2 S0 4 and salt in 
a lead-lined tank. This extracted the antimony and arsenic 
in the course of two or three days, when the antimony was 
to be precipitated out by diluting with water. The anti- 
mony dissolved, but the process had to be given up because 
no suitable apparatus for melting the insoluble portion of 
the slime was available. The melting was attempted in cru- 



CHEMISTRY OF SLIME TREATMENT. 127 

cibles, which were rapidly corroded by the basic fluxes used, 
and the capacity of the whole arrangement was too small. 
Also the cost of sulphuric acid and salt was quite a heavy 
item. Laboratory tests had showed an extraction of the 
antimony in about three days, and considerable confidence 
was unfortunately placed in the current statement in books 
that blowing air through slime suspended in sulphuric acid 
w T as an efficient means of oxidation. The long time required 
in the laboratory test was thought to be due to the small 
scale of operation and shallowness of the layer through which 
the air passed. This process is, however, available when 
salt and sulphuric acid are cheap and enough tank capacity 
is at hand. 

Air oxidation with sulphuric acid is probably consider- 
ably slower yet. The presence of iron salts, which are con- 
verted by air from the ferrous condition to ferric condition, 
might be thought to be an aid to the process, but the oxida- 
tion of acid ferrous sulphate solution for example, by air, 
is extremely slow. I thought at one time that if ferric sul- 
phate could be made in this w T ay and then used to attack 
slime it could be used over and over again, crystallizing out 
copper and arsenic occasionally and adding sulphuric acid 
to make up for that removed by copper. Various arrange- 
ments were tried unsuccessfully, including the use of platinum 
black as catalyzer. 

The necessary oxidation of the iron can, however, be 
easily secured in another way. The solution of ferrous sul- 
phate, resulting from the treatment of slime, which solution 
should be as hot and strong as possible, was cooled when 
ferrous sulphate crystallized out. The crystals were then 
gently dried and roasted, effecting a ready oxidation to basic 



128 LEAD REFINING BY ELECTROLYSIS. 

ferric sulphate, without loss of sulphur oxides. The product 
of basic ferric sulphate was completely soluble in the solu- 
tion. 

Roasting processes. — There are two classes of roasting 
processes for preparing slime for further treatment, one con- 
sisting in roasting the slime by itself, and one with the addi- 
tion of sulphuric acid.* Roasted by itself most slime ignites 
as soon as it is dry, large amounts of arsenic fume escaping 
and a yellow product resulting, which is largely unattacked 
by acicl solutions, even hydrofluoric acid. The antimony 
appears to be converted to a higher oxide, which resists all 
attempts to dissolve it, and the only further treatment avail- 
able is by melting. Starting with slime, however, previously 
rather well oxidized by drying or standing in the air, a more 
moderate reaction with air occurs and a less refractory pro- 
duct results. Even in this case the proportion of antimony 
soluble in HF as SbF 3 approximates only say 60%. The 
peroxidized antimony is not appreciably reducible by boiling 
with acid ferrous sulphate solution. 

Some slime can, however, be successfully oxidized by 
drying and heating at a moderate temperature, say 100-150°. 
The oxidation is not quite complete. The slime is next 
treated with hot dilute sulphuric acid and sodium nitrate 
added in sufficient quantity to complete the oxidation. Cop- 
per and arsenic are thereby extracted, and the residue after 
washing is leached with hydrofluoric acid for antimony-fluo- 
ride. The drying and heating is effected in long iron or lead 
pans heated by steam coils underneath for twenty-four to 
forty-eight hours. The slime is spread on in a layer about 4" 

* E. F. Kern, U. S. Patent 803,601. Nov. 7, 1905. 



CHEMISTRY OF SLIME TREATMENT. 129 

thick, in a lumpy condition, as removed from the filter. The 
extraction of the copper and arsenic is best effected in a lead- 
lined tank fitted with stirring gear. Filtration may be either 
done in a press, or on horizontal cloth filters with or with- 
out vacuum underneath. Quick filtration is the best, because 
no great cooling takes place, with consequent crystallization 
of arsenious acid or salts. On cooling the solution deposits 
a little antimony trioxide, and arsenious acid may crystallize 
out, if its concentration is high enough. 

The extraction of the antimony and treatment of the 
solution is the same as described on page 97. 

Roasting with sulphuric-acid process. — This is a simple, 
effective, and convenient method of oxidation. The first 
step is to mix the slime with concentrated sulphuric acid, 
which may be clone without drying the slime. The slime, 
however, will either be dry, or be in a cake from some kind 
of filter. The pasty or muddy mixture is then dried out on 
a plate or in a furnace with free air access. If sufficient sul- 
phuric acid is used to form lead, silver, copper, bismuth, and 
antimony sulphates, the product is mostly easily treated 
later, probably because the antimony sulphate, as soon as it 
touches water, decomposes and leaves a soft residue, whereas 
with less sulphuric acid present lumps are produced that 
are with difficulty completely attacked by the solutions. Dr. 
Valentine has suggested air-drying first, followed by roasting 
with sulphuric acid, as saving acid.* 

In either case acid fumes escape from the mixture on 
adding H2SO4, which are probably fumes of H 2 SiF 6 , or pos- 
sibly SiF 4 . The smell of the fume does not suggest HF. 

* Letter from Dr. Wm. Valentine. 



130 LEAD REFINING BY ELECTROLYSIS. 

During the heating the sulphuric acid carbonizes organic 
matter, a product of the glue added to the lead-depositing 
electrolyte, and in some cases has produced a product con- 
taining probably carbon in such form as to give the slime 
a greasy flotation. This is probably the result of the use of 
insufficient H 2 S0 4 . As a general average 1 lb. of slime will 
require T 6 o to | lb. sulphuric acid. 

A temperature of 200-250° C. for the roasting, which only 
takes about two hours with a layer f inch thick, is about 
right. As an excess of sulphuric acid is present the pro- 
duct is never a dry, dusty mass, and no dusting has ever been 
observed. The color of the product properly roasted is pur 
plish gray and it consists of silver, lead, copper, and anti- 
mony sulphates and gold and sulphuric acid. In what con- 
dition the arsenic is is not known, but it is probably AS2O3. 
Some small amount of arsenic is probably volatilized, but 
the quantity lost is certainly small. 

The destruction of gelatine left from the lead-depositing 
solution by the hot sulphuric acid is an advantage, for the 
resulting solutions settle and filter with greater ease than 
is the case with other wet methods. 

The product need not be ground if sufficient sulphuric 
acid has been used. It is boiled up with water, using suffi- 
cient to dissolve the arsenic present. For this prupose not 
less than 15 parts water should be added for each part of 
arsenic known to be present. Considerable silver dissolves, 
but only from one-third to one-half of the total, so no 
attempt is made to separate the silver, but copper is sus- 
pended in the hot mixture until silver has been removed 
from the solution. In the filtrate is practically all the cop- 
per, 80 to 90% of the arsenic, and about 2.5 grams antimony^ 



CHEMISTRY OF SLIME TREATMENT. 131 

and 2 grams bismuth per litre, if bismuth is present in the 
slime. 

Several methods of treating the solution for arsenic, cop- 
per, and bismuth may be adopted, as crystallization for cop- 
per sulphate and arsenious acid, precipitation of copper by 
scrap iron or electrolysis of the solution with a lead anode 
for electrolytic copper, and an arsenious solution, from which 
As 2 3 , mixed with some Sb 2 3 , may be crystallized. The 
AS2O3 may be further refined by sublimation or by crystalli- 
zation from hot water, to which a little HF is added to keep 
antimony in solution. 

The first method will probably not easily yield copper 
sulphate free from arsenic, and I have not attempted it. The 
second method has been in use in practical work, but the 
copper only comes down slowly on scrap iron, the process is 
wasteful of iron and acid, and the product is a low-grade one. 
With the third method the cost of electrolytic precipitation 
as pure copper is less, the product is a finished one, and there 
is no loss of acid, and the separation from arsenic is nearly 
complete. The sulphuric acid may also be used over again, 
after concentration. 

In one experiment the conditions were as follows: Cop- 
per volume percentage on start 3.5%, on finish 0.53%. 
Cathode current density 18 to 9 amperes, and even as low 
as 4.5 for a short time. Volts 2.3 to 2.1. Copper deposited 
at finish, good color. Anode of lead from one-quarter size of 
cathode most of the time, to same size as cathode. Current 
efficiency approximately 100%, but not accurately deter- 
mined. A good agitation was maintained, but about the 
middle of the run, with current density 20 amperes, deposit 
got black on top for a short time. 



132 



LEAD REFINING BY ELECTROLYSIS. 



The solution from which copper was removed was evapo- 
ated clown until As 2 3 began to crystallize out, when arsenic 
was mostly removed in hard, glittering crystals intermixed 
with Sb 2 3 and some copper sulphate, which was washed out 
with water. 

A better procedure might be to cool the hot acid solution 
filtered from the slime, crystallize out copper sulphate and 
arsenious acid, and dissolve the copper sulphate from the 
product with water or with similar solution from a previous 
treatment from which the copper has been largely removed 
by electrolysis, leaving the crude arsenic insoluble. Any bis- 
muth present in the hot solution from the filter, remains in 
solution, as bismuth is as soluble or more soluble cold than 
hot, except in very strong acid. On evaporating the mother 
liquor down for crude sulphuric acid for the treatment of 
another lot of slime, bismuth and remaining copper sul- 
phate mostly separate, or can be separated by cooling the 
strong sulphuric acid. 

The solubility of bismuth sulphate in sulphuric acid of 
various strengths is approximately as given in Table 49, from 
experiments by Dr. E. F. Kern: 





TABLE 49. 


97% H^O, 


.69 grai 


55% H^O* 


.61 " 


16.6% H-SO* 


.19 " 


10% H^O, 


.22 " 


10% KSO, 


.21 " 


3% ELSO, 


.16 " 


4% H^O, 


.10 " 



69 grams Bi per 100 cc. cold. 







' 100 


I i 








' 100 


I { 








' 100 


i i 








1 100 


c c 








-—1 


( I 








' 100 


I c 





In general bismuth is more soluble cold than hot in 
weaker solutions. Thus with 20% H 2 S0 4 the solubility is 



CHKMISTRY OF SLIME TREATMENT. 



133 



greater cold; with 50% II_>S0, and stronger acids, the solu- 



bility is greater hot. 




32% H 2 S0 4 


100° c. 


88% " 


100° c. 



.65 gr. Bi per 100 cc 



1.86 



100 



With relatively very small amounts of bismuth present 
most of it will be removed from the slime with the copper 
and arsenic. With large amounts of bismuth most will 
remain in the slime throughout. 

The extraction with HF for antimony proceeds as de- 
scribed under heading " Ferric Sulphate Process", page 97, 
but gives even cleaner extraction in this case. Some silver 
dissolves, which is readily precipitated out with metallic anti- 
mony. The following figures are for Trail slime treated 
experimentally. The figures in the second column are not 
exactly right, probably partly on account of absorption of 
moisture since the analysis was made. 

TABLE 50. 



Silver .... 
Gold .... 
Copper. . . 

Lead 

Arsenic. . . 
Antimony- 
Bismuth. . 



Slim? 
100 Grams. 



First Residue 
74 Grams. 



14.6% 
34.5 ozs. 

8.1% 
16% 

7.0% 
27.60% 

0.81% 



12.3% 

Tr. ' 

1.66% 
35.9% 



Second Residue 
33.8 Grams. 



20.9% 

None. 

None. 
0.56% 



The amounts of metals in the various products are given 
below. The discrepancy in silver is due to the fact that in 
this experiment the silver, instead of being precipitated back 
into the slime as would be done in practice, was precipitated 



134 



LEAD REFINING BY ELECTROLYSIS. 



separately in the filtrate in order to determine the quantity 
in solution. 



TABLE 51. 



In Slime. 


In First Residue. 


In Second Residue. 


14.6 gr. silver 
8.1 gr. copper 

16.0 gr. lead 
7.0 gr. arsenic 

27 . 6 gr. antimony 


9.1 gr. silver 
No copper 
16 gr. lead 

1 . 23 gr. arsenic 
26.6 gr. antimony 


7.1 gr. silver 
No copper. 
16 gr. lead. 
No arsenic. 

. 19 gr. antimony. 



The extraction of copper by the first solution was about 
100%, of arsenic 80%, of antimony (from other facts), 2.6%. 
The extraction of arsenic by the HF is the remaining 20% 
of the arsenic, and of the antimony about 96.6% of the total 
originally present. The result in respect to antimony is 
superior to that obtained by the ferric sulphate method on 
dried slime. 

In another experiment with 600 gr. lots of the same slime. 
the following results are given, arranged as Table 52 (p. 135) 
for the sake of brevity. 

Dissolving air-dried slime in H 2 SiF 6 and HF. — Practi- 
cally all the slime, if sufficiently well oxidized, dissolves in 
a solution containing both fluosilicic acid, and a moderate 
quantity of hydrofluoric acid, in a few hours. Lumps disin- 
tegrate of themselves. The solution resulting contains lead 
and copper fluosilicates, antimony fluoride, and arsenious 
acid. For the recovery of the various metals experimentally, 
the solution was electrolyzed first with an antimony anode 
and copper cathode, current density 2 amperes per square 
foot, e.m.f. .25 volts. Good copper comes down, if the solu- 
tion is stirred until copper is nearly all gone, when the deposit 



CHEMISTRY OF SLIME TREATMENT. 135 

TABLE 52. 



Experiment 



Experiment 2. 



Slime taken 

H.S(\ taken calculated to H\,S0 4 

Sulphuric acid lost in roasting percent- 
age of slime taken 

Copper removed from solution 

Anode 

Copper in solution on start 

Copper in solution on finish 

Quality of copper with hot solution 

Water used in dissolving sulphates 

Amperes per sq. ft. in copper deposition 
maximum 

Amperes per sq. ft. in copper deposition 
minimum 

Volts, maximum 

minimum 

CuS0 4 crystallized from mother liquor. . . . 

Copper dissolved in precipitating silver. . . 

Wt. dry residue after dissolving sulphates 

Antimony dissolved to remove Ag from 
fluoride solution 

Antimony deposited 

Quality 



Arsenic in electrolyte 

Amperes per sq. ft. cathode surface, maxi- 
mum 

Amperes per sq. ft. cathode surface, mini- 
mum 

Amperes per sq. ft. cathode surface at end 
of electrolysis 

Volts at 20 amperes per sq. ft 

Current efficiency 

Antimony as trifiuoride in solution on 
start, grams per litre 

Antimony as trifiuoride in solution on 
finish, grams per litre 

Antimony as pentafluoride on finish, 
grams per litre 

Free HF in solution on start, grams per 
litre 

Antimony loss in whole of two operations . 

Insoluble residue melted with silica, giving 
clean dore weight 



600 gr. 
450 gr. 

33% 
By electrolysis 
Lead 

24 gr. per litre 
4.0gr. " " 
Varied 

5 times weight 
of slime 

26 

9.4 

2.45 

2.0 

18.4 gr. + 

24 . 5 gr. 
462 gr. 

7 gr. 
114 gr. 
Fair, contained 
copper not 
thoroughly 
washed out 
before treat- 
ment with HF 
0.3% 

31.5 

15.6 



600 gr. 
400 gr. 

22% 
By electrolysis 
Lead 

21 gr. per litre 
2.3gr. " " 
Good 

5 times weight of 
slime 

9.3 

2.2 



16 gr. 
15 gr. 



12 gr. 
149 gr. 
Excellent , 
copper 



0.6% 
24 



4.2 



16.5 

2.78 
84.5 


11.8 
3.05 
97.5 


91.4 


108 


7.6 


6.9 




18.9 


25.0 


14.0 
7.5% 


88 gr. 





136 LEAD REFINING BY ELECTROLYSIS. 

turns whitish. Analysis of copper product, 91.1% Cu; 4.8% 
Sb; .25% Bi; no As or Pb. The next product with copper 
cathode and antimony anode is small in amount and con- 
sists of antimony with about 10% copper. 

The solution is next electrolyzed with a lead anode and 
copper cathode, current density 10 amperes per square foot, 
e.m.f. 0.2 to 0.4 volts. The antimony deposit contained 
90.5% Sb, 5.6% As, no Cu. Some antimony also forms on 
the anode, as scale. 

The current density should be diminished when antimony 
is reduced to 2%, to prevent lead from coming down. 

The solution is then electrolyzed with lead anode and 
cathode. A soft deposit forms on the cathode, which can be 
compressed to solid metal. Analyses show, for successive 
products : 

TABLE 53. 



Pb 




86.9% 




N. d. 


Sb 


7.2% 


9.2% 


4.4% 


Trace. 


As 


•6% 


3.3% 


19.6% 


n 



By reversing the current an anode slime of arsenic and 
antimony may be produced. 

The solution is next electrolyzed with carbon anode and 
lead cathode for the production of PbC>2 and Pb, containing 
arsenic and antimony and free acid to be used over again, 
in the treatment of another lot of slime. The operation some- 
times succeeds and sometimes no Pb02 separates at the 
anode, for reasons not understood. 

The solution at different times contained approximately 
as follows: 

Column 1 shows the solution after filtering from slime, 
column 2 after removal of copper, column 3 after removal of 



CHEMISTRY OF SLIME TREATMENT. 137 

antimony, column 4 after removing arsenic and remaining 
antimony in lead, and column 5 after electrolysis for Pb0 2 
and Pb. The Pb0 2 ean be put with a charge of lead ore for 

recovery of lead. 

TABLE 54. 



Cu" 


1.3% 


% 


% 


0% 


% 


Sb" 


o % 


6.6% 


44% 


Trace 


% 


Pb" 


5.5% 


5.5% 


21.5 % 


21.5% 


5 % 


SiF„" 


17 % 


17 % 


17 % 


17 % 


17 % 


P 


3.5% 


3.5% 


3.5 % 


3.5% 


3.5% 


As'" 


1-1% 


1-1% 


• 88% 


Trace 


% 



Analysis of 50 parts of air-dried slime before treatment 
and 13.5 parts residue after treatment with the mixed elec- 
trolytes, Table oo. 

TABLE 55. 

Ag 14.6% 

Cu 8.1% 7% 

Pb 16.0% 7.2% 

Sb 27.6% 12.5% 

As 7.0% 1.6% 

Au 

The analysis of the slime taken is not exact as it had 
probably absorbed water, and consequently the values may 
be a trifle too high. 

The percentage of extraction was as follows, Table 56: 

TABLE 56. 

Copper 73% extracted. 

Lead 86.2% 

Antimony 87. 7% 

Arsenic 93.8% 

Silver None 

Gold " 



CHAPTER III. 

DEPOSITION OF ANTIMONY FROM THE FLUORIDE SOLUTION. 

The electrolytic refining of antimony with an electrolyte 
containing SbF 3 and HF, and perhaps KF or NaF in addition,* 
is a successful method, as far as the quiet solution of the anode 
and good mechanical quality of the cathode is concerned. 
The addition of KF or NaF is made to increase the conduc- 
tivity of the solution. Dilute hydrofluoric acid is not as good 
a conductor as the other common acids, sulphuric acid for 
example. KF also removes H 2 SiF 6 from the solution as a 
precipitate of K 2 SiF 6 . It has been proved* that the presence 
of sulphuric acid or sulphates in the refining electrolyte (to 
be distinguished from the electrolyte when insoluble anodes 
are used, as described later) prevents the easy solution of the 
anodes, which is readily explained by the ionic electrochemical 
theory, as follows: The anion S0 4 , with a smaller quantity 
of the anions F' or F" 2 whichever is produced b}^ the disso- 
ciation of HF, combine with the anode metal to form anti- 
mony sulphate and antimony fluoride. Antimony sulphate 
is almost immediately decomposed into insoluble antimony 
oxide or hydroxide and sulphuric acid, thus leaving an insu- 
lating coating on the anode, which is only slowly, and in fact 
too slowly, dissolved off by free HF which may be present. 



* Betts Trans. Am. Electrochemical Society, Vol. VIII, page 190. 

138 



DEPOSITION OF ANTIMONY FLUORIDE SOLUTION. 139 

This insulating coating actually exists, and may produce a 
local resistance sufficient to absorb .2 volt or more. As the 
difference of e.m.f. of solution of copper and antimony in 
the fluoride solution is probably considerably less than .1 
volt, only a slight voltage drop is necessary to make any cop- 
per present dissolve too, and once dissolved, it readily de- 
posits on the cathode with the antimony. 

Whether copper can be left as anode slime, in absence of 
H0SO4, H 2 SiF 6 , etc., and antimony free from copper can be 
produced in this way, has not been definitely settled. 

Arsenic probably dissolves even more readily than anti- 
mony and collects in the solution, though some will be found 
in the cathode metal under some conditions, if not all. 

Lead is eliminated satisfactorily, provided suitable cath- 
odes of other material than lead are used. 

Antimony trifluoride is an extremely soluble salt. Its 
cold saturated solution in water has a specific gravity of about 
2.6 and contains about three parts SbF 3 to 1 part H 2 0. By 
adding other salts as sodium, ammonium, potassium, chlo- 
rides, fluorides and sulphates, double salts of less solubility 
may be secured. Antimony trifluoride is used as a mordant 
in dyeing, though probably better results are got with anti- 
mony lactate and tartar emetic. 

The deposition of antimony from the trifluoride solution, 
which in this case may well contain H 2 S0 4 or sulphates, is 
important in working up anode slime, as it is often conven- 
ient to dissolve the antimony oxide in oxidized slime, or slags 
from melting slime in dilute hydrofluoric acid, followed by 
deposition of the antimony from the solution. With an in- 
soluble anode the principal reaction is 



140 LEAD REFINING BY ELECTROLYSIS. 

(1) 2SbF 3 +3H 2 O = 2Sb+30+6HF. 

There is a secondary reaction that takes place, namely, 

(2) 5SbF 3 = 2Sb+3SF 5 . 

The last reaction is undesirable, as the antimony in SbF 5 
represents a loss of both antimony and fluorine as the process 
is worked at present. It is hoped to devise means to reduce 
this SbF 5 again to SbF 3 , but no serious attempt has been made 
yet. 

Reaction 2 is favored by high percentage of SbF 3 , high 
temperature, low percentage of H 2 S0 4 , large anode surface, 
and ready access of solution to the anode surface, so the oppo- 
site conditions are adhered to in practice, when reaction 2 
may be reduced to about 5% of the whole electrochemical 
effect. 

The , available anode materials are platinum, carbon, and 
lead. It is quite possible that fine platinum wires would 
make an excellent and permanent anode, but they have not 
been tried. Carbon anodes of all kinds distintergrate rapidly, 
and can only be used when the solution is supplied with 
some reducing agent, as S0 2 . This is of course converted 
at the anodes into H 2 S0 4 , and might be used practically, 
except that it is also reduced at the cathode, forming Sb 2 S 3 
Lead anodes only are actually used, but it is necessary to use 
them in a special manner, both to save lead, and to prevent 
the formation of much SbF 5 . 

The commercial hydrofluoric acid used in extracting an- 
timony from slime and slags from melting slime, contains 
H 2 SiF 6 , and as the slime or slag usually contains silica, further 
quantities of H 2 SiF 6 are formed. Dr. Wm. Valentine has 
noticed that in making HF by distilling fluorspar with sul- 



DEPOSITION OF ANTIMONY FLUORIDE SOLUTION. HI 



phuric acid, the first HF to come off contains most or all of 
the silica, and has suggested using the first part in making 
lead refining electrolyte and the last in slime treating. 

The presence of fluosilicic acid (or any acid forming a sol- 
uble lead salt) is undesirable, for it acts on the lead anodes 
as a strong "forming" agent, and therefore reduces the life 
of the anodes. H 2 SiF 6 is usually removed sufficiently by 
precipitation with sodium sulphate. Potassium sulphate 
is better, but its cost has been too high. However, as the 

ft 




sodium fluosilicate is too valuable to throw away, and should 
be distilled with H2SO4 and a little fluorspar anyway, to re- 
cover the H 2 SiF 6 , potassium sulphate would be equally as 
economical, for the residual potassium sulphate could be used 
over again. 

To test the anode reactions, antimony trifluoride solution 
containing also ferrous sulphate, to imitate conditions in prac- 
tice when iron gradually accumulates in the solution, was elec- 
trolyzed in series with a gas voltameter, provisions being made 
for collecting the gas liberated at the anode. 

Apparatus as shown in Fig. 21a was used. A is a gas 



142 



LEAD REFINING BY ELECTROLYSIS. 



voltameter using lead anode and cathode in an acid solution 
of copper sulphate. B is a resistance cell for regulating the 
current. C contains the electrolyte under investigation and 
has a small lead anode from which the escaping gas can 
be collected and measured in the burette. The results are 
tabulated in Table 57. 



TABLE 57. 





Amperes 




Percentage of 
Current Used 




No. 


Square 
Foot 


Average 


Solution. 




Voltage. 


in Generating 






of Anode. 




Oxygen Gas. 




1 


87 


3.4 


75.3 




2 


• 87 


3.4 


77 


7 . 5 gr. SbF 3 1 per 


3 


85 


3.3 


81.6 


5 gr. HoS0 4 100 


4 


86 


3.3 


82.4 


25 gr. FeS0 4 -7H 2 J cc. 


5 


88 


3.2 


71.5 




6 


81 


3.2 


72.7 




7 


87 


3.1 


72.6 




8 


85 


3.3 


71.4 


15 gr. SbF, 


per 


9 


75 


3.2 


71.7 


40 gr. FeS0 4 -7H 2 


.100 


10 


73 


3.1 


66.5 


5gr. H 2 S0 4 ' J 


cc. 


11 


65 


2.9 


65.1 




12 


45 


2.6 


52 




13 


31 


2.7 


42.5 





The figures given for percentage of current used in gen- 
erating oxygen gas give, by subtraction from 100, the per- 
centage of the current used in oxidizing ingredients of the so- 
lution, which is not desired. If iron is oxidized the ferric 
salt will react at the cathode and cut down the efficiency, 
and any antimony oxidized results in temporary loss of anti- 
mony. 

The highest efficiency in Table 57 is surpassed in prac- 
tical work with lead rods as anode, wrapped in several thick- 
nesses of cloth to prevent the free access of oxidizable salts 
to the anode surface. 



DEPOSITION OF ANTIMONY FLl/ORl!)!-; SOLUTION. 



143 



The accompanying Table 5S shows the efficiency in ex- 
periments in depositing antimony where the efficiency was 
accurately determined anil other data carefully noted. 

In experiment 2 in the table, no cloth was wrapped around 
the anode rods and the lower efficiency should be noted. 

TABLE 58. 











Cathode 


No. 


jj ate Quantity 


Current measured | Current Density Current Density 




Deposited. 


b y per Square Foot, per .Square Foot. 


1 


March 1905 25.7 gr. 


Lead voltameter 375-60 amps. 25 . 4 amps. 


2 


Sept, 1903 1.077 kg. 


Ammeter 180-120 22.5-15 " 


3 


Sept. 1903 .91 " 


" 


75-38 


19-9.5 


4 


Oct. 1903 .S74 '• 


" 


100-40 


25-10 


5 


Oct. 1903 1.243 " 


< t 


92-52 


23-14 


6 


Oct. 1903 1.585 " 


l c 


105-34 


26-8 .5 


7 


March 1907 114 gr. 


" 


102-51 


31.5-15.6 • 


8 


March 1907 149 " 


I i 


120-21 


24-4 .2 









Sb'" on Start 


Sb'" on Finish 




No. 


Date. 


Efficiency. 


in Solution 
per 100 cc. 


in Solution 
per 100 cc. 


Volts. 


1 


March 1905. 


90.0% 


3 . 48 gr. 


0.80 gr. 


3-2S 


2 


Sept. 1903' 


66.5 


6.8 " 


1.36 " 


3-2 . 75 


3 


Sept. 1903 


84.5 


5.52 " 


1.83 " 


2.9-2.7 


4 


Oct. 1903 ! 


92.9 


5.58 " 


1.83 " 


3.15-2.45 


5 


Oct. 1903, 


95.4 


7.33 " 


2.07 " 


2.9-2.55 


6 


Oct. 1903! 


92.0 


8.1 " 


2.15 " 


2.9-2.55 


7 


March 1907 


84.5 


9.14 " 


.76 " 


3.05-2.78 


8 


March 1907, 


97.5 


10.8 " 


.69 " 





No. 



Date. 



'Sb'"" on Fin- 
nish in Solution 
per 100 cc. 



NasSO^ per 

100 cc. 



H2SO4 per 
100 cc. 



HzSiFs per 

100 cc. 



'F' pc;r 100 cc, 



March 1905 

Sept. 1903| 

Sept. 1903! 

Oct. 1903! 

Oct, 1903 

Oct. 1903 

March 1907 

March 1907 



1.89* gr. 



1 . 5 gr. 
.24 gr. 

"ASgr. 

Excess 



4 . 66 gr. 
l.Tgr." 
3.'o'gr." 



0.8 gr. 
1. gr. 



4 


1 gr 


4 


1 " 


4 


4 " 


3 


2 ": 


4 


5 " 


3 


87 " 


7.0 


6 


45 " 



* Total amount produced in runs 7 and 8 = approx. 10%. 



144 LEAD REFINING BY ELECTROLYSIS. 

The anodes are of soft lead rods, usually J to §" diameter, 
and covered with 2 to 4 layers of cotton cloth to prevent the 
access of much SbF 3 to the actual anode surface with its oxidi- 
zing conditions. Oxygen escapes vigorously while the current 
is on. The anode rods are spaced about 3 inches apart in 
rows with cathode plates between. An experimental tank 
is described and illustrated on page 396, and a commercial 
tank on page 260. 

The electrolytic antimony may be pure or not, according 
to the solution used. When lead cathodes are used the an- 
timony is found to contain lead, the reason being evident 
to anyone who examines the corrosion of a lead cathode, when 
such has been used. A copper cathode is more satisfac- 
tory. 

The presence of lead in the antimony is avoidable and 
so is that of copper. If the solution contains copper, it comes 
down with the first antimony deposited, it being necessary 
to deposit perhaps one-tenth of the total antimony to get 
the copper all out. In practical work, however, little or no 
copper is found in the solution anyway. 

The removal of the copper is better carried out before 
the electrolysis in either one of two methods, or if the quan- 
tity is large, by a combination of the two. For considerable 
quantities of copper the solution is electrolyzed with antimony 
chunks as anode and copper cathodes. With a cathode cur- 
rent density of 2 amperes per square foot and anode current 
density, which may be as high as 10 amperes and .4 to .5 volts 
practically all the copper can be got out as good copper, while 
of course a corresponding amount of antimony goes into solu- 
tion. For results with this method see Table 59. 



DEPOSITION OF ANTIMONY FLUORIDE SOLUTION. 145 
TABLE 59. 



No. 


Copper on 
Start 


On Finish. 


Anode CD. 
uor Sq. Ft . 


Cathode per 
Sq. Ft. 


Remarks. 


1 
2 
3 


•24% 

• 50% 

1.00% 


Trace 
.03 


About 10 

2-8 
About 5 


2 

2-8 
3-2 


Contained much 
H 2 SiF 6 



Small quantities of copper may be conveniently removed 
by direct precipitation on antimony. While merely drop- 
ping some antimony into a tank containing the coppery solu- 
tion does little or no good, an arrangement as shown in Fig. 
216 is successful, especially if the solution passes through slowly 




Fig. 216. 

and at a slightly raised temperature, say 40° C. The tank 
contains broken antimony resting on a false bottom, in a layer 
4" or more thick. Copper deposits on the top of the mass, 
while antimony dissolves away underneath. The solution 
escaping has a yellowish color and probably contains traces 
of copper as cuprous fluoride. 

The removal of arsenic is not readily accomplished be- 
fore the electrolysis, and the best way seems to be to let it 



146 



LEAD REFINING BY ELECTROLYSIS. 



accumulate in the solution, which it may be expected to do 
in practice at the rate of about 1 part arsenic or less dissolved 
for 30 parts antimony deposited (see page 98, Chapter II). 
Antimony deposits more readily than arsenic. 

For analyses of electrolytic antimony from slime, see 
Table 60. 



TABLE 60. 
Analyses of Antimony. 



No. 


Ag 


Pb 


Cu 


As 


Bi 


Sb 


1 


Nil 


1.6% 


2.9% 


2.3% 




93.8% 


2 


i i 


.62% 


•2% 




.67% 




3 


t i 


Nil 


•07% 


•41% 


Nil 




4 


1 1 


Nil 


Trace 


0.5-1.00% 


Nil 





No. 1 from solution not purified from copper used over 
and over in consecutive treatments and deposited on lead 
cathodes. 

No. 2 from slime containing much Bi. No refining of 
solution from copper necessary in this case. 

No. 3 from solution purified of copper before electrolysis. 
Deposited on copper cathodes. 

No. 4 from commercial work. Poorer quality also pro- 
duced. Arsenic is hard to keep down. 

In general, the antimony deposited will contain 0.5 to 
1.0% arsenic, and no easy method is known so far of produ- 
cing antimony free from arsenic in this way. However, ar- 
senic is about the easiest to remove in the dry way of the 
metals we consider here. Arsenic in the presence of bases 
is more oxidizable than antimony and can be slagged off as 
sodium arsenate by fusion under soda in presence of oxidi- 
zing agents. 



DEPOSITION ov ANTIMONY FLUORIDE SOLUTION. 147 

The deposited antimony is usually solid and hard, with a 
jagged but bright surface. However, when antimony becomes 
reduced to from 1 to 2%, according to the current density, 
the deposit gets black and soft, probably due to arsenic com- 
ing down too, and it begins to fall from the cathodes as powdc r. 
At this point, the operation should be stopped. The de- 
posited antimony shows a tendency to peel, but does not usually 
fall off the cathodes. It is easily removed from the cathodes, 
particularly as these are flexible and the brittle antimony 
separates readily on bending. 

The deposit appears to contain some of the solution, as 
acid fumes escape on melting, and the metal loses slightly 
in weight. 

As lead slime usually contains excess of silica, averaging 
perhaps 1-2% silica in addition to some fluosilicic acid or fluo- 
silicate, when antimony is extracted with HF, some silica 
also dissolves, varying in amount from 0.9 to 1.8% calculated 
on original w T eight of slime. By precipitation with sodium 
or potassium sulphate sodium fluosilicate is produced, which 
can be utilized by adding it in with a charge of fluorspar in 
the hydrofluoric acid plant. 

Cost of depositing antimony from the fluoride solution 
with insoluble anodes. — This process has not been used on 
a practical scale long enough for actual operating costs to 
be determined, but the cost can be quite closely estimated. 
The tanks for practical work may take 4000 amperes, at 2.8 
to 3.0 volts, and are 7 feet 2 inches long, 2 feet 6 inches wide, 
and 3 feet 6 inches deep. Current, 15 amperes per square foot 
of cathode surface. Anode, 20 sets of 10 lead rods, each §" 
diameter. 



148 LEAD REFINING BY ELECTROLYSIS. 

TABLE 61. 

Per Pound 
Antimony. 

Power cost at $50 per E.H.P. year at 95% efficiency, 1.20 H.P. 

hours at $0.00575 $0.0069 

Breaking antimony from cathodes and labor cost operating tank. . 0.0010 

Melting antimony in crucibles and casting . 0010 

HF loss, mechanical, 5% . 0018 

HF loss from formation of H 2 SiF 6 0.0057 

Na 2 S0 4 , .12 lbs. at $15 per ton 0. 0009 

Labor cost, precipitating and collecting Na 2 SiF 6 . . 0010 

Renewals of anodes (in 7 days 3 lbs. Sb deposited per ft. anode used), 
Smelting and refining and squirting 0.18 lbs. lead at $20 per ton 

including losses . 0018 

Cloth and labor wrapping anodes. . 0020 

Repairs and interest . 0020 

Total $0.0241 

Credit for Na 2 SiF 6 , added to fluorspar in making HF, yield H;>SiF 6 = 

80% . 0040 

Net cost per lb. antimony deposited $0.0201 



CHAPTER IV. 

ELECTROLYTIC REFINING OF DORE BULLION. 

The older nitric-acid and sulphuric-acid processes are well 
described in various works on the metallurgy of silver and 
gold,* to which the reader is referred. 

The electrolytic processes are now coming largely into 
use, and it is doubtful if the sulphuric-acid process will be 
much installed in future in large works. 

Further improvements in the electrolytic processes may 
be expected, particularly for alloys containing copper, so that 
the sulphuric-acicl process will fall farther behind than ever. 
In the older parting processes it was desired to remove the 
base metals as fully as possible to save acids in parting and 
make the process more easily conducted. This was done 
at quite high cost, and not without losses, by cupellation and 
furnace treatment, and the practice is still in vogue even at 
plants using the electrolytic processes, because in that way 
there is required less renewals of the electrolytes to get rid 
of the accumulating base metals and keep the silver percent- 
age at the necessary amount. However, the accumulation 
of base metals in the electrolyte need not necessarily be a 
disadvantage and in the future it will be found better to leave 
out the furnace refining for bullion from anode slimes, and 



*Rose, "Metallurgy of Gold"; Eissler^ "Metallurgy of Gold." 

149 



150 LEAD REFINING BY ELECTROLYSIS. 

recover the base metals present from the electrolyte 
instead, which it is easy to do in many ways, and will permit 
greater economy than long and expensive furnacing with 
unavoidable metal losses. 

The parting process of Dr. Dietzel* is based on sound 
principles, which ought yet to be more largely applied in 
better apparatus. Alloys of gold, silver and copper, contain- 
ing 

Au 5-7% Zn, Sn, Pb, about 5% 

Ag 22-50% Cd, Fe, Ni, Pt, Traces 

Cu 40-65% 

were successfully treated on a rather small scale. 

The process consists in electrolyzing a solution of copper 
nitrate with copper cathode and bullion anode, separated 
by a diaphragm, copper depositing on the cathode of course, 
and all the metals except gold dissolving from the anode. 
With alloys containing 40% silver or less I have found it 
difficult to dissolve any silver from hanging electrodes, as the 
other metals dissolve first and leave the silver as a mushy 
anode slime, and the same objection was probably found by 
Dr. Dietzel, as his apparatus has a horizontal conducting anode, 
of carbon probably, on which the alloy rests, and in this way 
of course the silver may be finally dissolved. 

There is maintained a continuous flow of copper nitrate 
solution to the catholyte, while the anolyte containing silver 
overflows and runs to a precipitating tank in which the sil- 
ver is cemented out by copper, and the solution then goes 
back to the electrolytic cell. This system has important 
advantages which should not be lost sight of. 

* Borchers, "Electric Smelting and Refining," 2d Eng. Ed., page 304. 



ELECTROLYTE REFINING OF DORK BULLION. 151 

(1) The system is perfectly cyclic (except if the anodes 
contain iron, zinc, tin, ami lead), and so little or no mainte- 
nance of solution is required. 

(2) The process is not affected by variation in the compo- 
sition of the anodes, as it will work the same on a series of 
alloys all the way from pure copper on one end to pure silver 
on the other. 

(3) All the silver is precipitated in one or two tanks, 
and the superiority of this plan over collecting spongy silver 
from a large number of different cells is apparent. 

(4) Whatever copper is present in the anodes appears 
as electrolytic copper. 

One objection, though not* readily apparent, may be noted. 
If selenium or tellurium, or other metal precipitable by copper 
dissolves from the anodes the silver will contain that element. 
This objection applies to most precipitation processes. Whether 
by a partial precipitation of the silver the selenium or tellu- 
rium could be concentrated in a small part of the silver, is not 
known, but it seems that this probably could be done. Under 
some quite usual conditions I clo not believe, however, that 
selenium or tellurium would dissolve with the silver. These 
conditions are found when the anode contains a preponderat- 
ing amount of silver. 

With the Dietzel process it will be seen that the silver 
and copper solution escaping from the anode compartment 
might be strong enough in silver to provide electrolyte for 
an electrolytic silver refining cell, while the electrolyte from 
the latter, impoverished in silver, might then be carried 
through the rest of the process as originally intended. For 
an ordinary Moebius or analagous parting plant, the use of a 
number of cells on the Dietzel principle would be desirable, 



152 LEAD REFINING BY ELECTROLYSIS. 

as providing a means of recovering copper from and return- 
ing silver to the electrolyte. 

The cell used by Dr. Dietzel does not appear to be espe- 
cially well suited to the work, however. The use of rolling 
cylinders of copper cathodes would seem unnecessary. For 
most dore bullion, the percentage of silver is so high that 
the silver may be cast directly to anodes and suspended or 
supported in the solution, instead of requiring a flat surface 
on the bottom on which to support the pieces of dore anc. 
the slime, still containing considerable silver in that case. 

A diaphragm cell with diaphragms of porous earthenware 
or asbestos sheets supported between perforated slate or glass 
plates, or absestos plugs in holes in a wood partition, or one 
of hardened asbestos (see page 110) can all be expected to 
give a good result, of which the one objection is that silver 
moves under the action of the current through the diaphragm 
and toward the cathode, and interferes with the deposition 
of a solid smooth cathode. This objection (not a very serious 
one) can be got around by using a double diaphragm and in 
the space between a piece of metal, for example copper, to 
precipitate silver. I have tried this arrangement, but the 
results are not conclusive either way. 

For refining bullion containing lead and bismuth appar- 
atus as shown in Fig. 22 gives good results. The bullion 
is placed as anode in cell 1 with a lead cathode and a diaphragm 
of sulphurized asbestos between. A steady flow of lead methyl 
sulphate solution containing 5-6% Pb and 12-15% CH 3 S0 4 , 
is maintained to the cathode compartment in which lead 
is deposited in a fair condition of solidity. This lead is not 
pure, however, and in practice would go to the lead-bullion 
kettle. At the dore anode, silver, copper, bismuth, and lead 



ELECTROLYTIC REFINING OF DORE BULLION. 



153 



dissolve, provided the dore contains approximately 70% 
silver or over. If less bismuth and lead dissolve and leave 
a mushy anode slime of silver containing about 15% of lead 
and bismuth. The solution continually overflows from the 
cathode compartment where the percentage of lead is reduced, 
to the anode compartment, while solution containing silver, 
lead, bismuth, and traces of copper flows through a series of 
beakers to a storage vessel. The first tw r o contain pieces of 




Fig. 22. 



bismuth, which cement the silver out readily, and the last 
two contain metallic lead, which throws out the bismuth. 
The solution, practically free from copper and bismuth, is ele- 
vated to the higher storage tank and passes through the series 
again. The current density in the electrolytic cell was 15 
amperes per square foot. Electromotive force, 2 volts. 

A process for getting the silver into solution quickly at 
the anode, without introducing any difficulties at the cathode 



154 LEAD REFINING BY ELECTROLYSIS. 

in the way of producing a solid deposit of silver, or lead, or cop- 
per, as the case may be, is furnished by the use of lead perox- 
ide and a solution of fluosilicic acid, for instance. The dore 
may be dissolved at a very high current density if the lead 
peroxide is used as cathode, especially if the cathode is of 
carbon electrolytically coated with the peroxide. There 
results a solution of lead, copper and silver fluosilicates that 
may be rapidly treated for silver by precipitation on copper, 
while the copper can be got out by electrolysis with lead anode 
and copper cathode, and next the lead can be removed and 
the lead peroxide used recovered by electrolysis with carbon 
anode and lead cathode; or equally well, if the dore contains 
little lead, the precipitation of copper as mentioned above 
may be omitted, and the solution containing copper and lead 
fluosilicates can be electrolyzed with a carbon anode and lead 
cathode for electrolytic copper and lead peroxide, the latter 
of course being used over again as cathode in dissolving more 
bullion. 

In this process there is no difficulty either with the cathode 
deposits being spongy, or is there need to consider the dia- 
phragm question, but another difficulty appears in that the lead 
peroxide deposits on carbon electrodes have not yet been 
dissolved off with high efficiency, some of the peroxide drop- 
ping from the electrodes to the bottom of the cell and thereby 
escaping action. To obviate this, a plate of bullion or a plate 
of graphite connected electronically to the peroxide cathodes 
might be placed on the bottom of the cell. The peroxide fall- 
ing on the bottom in this case would be ultimately reduced 
and dissolved, though somewhat slowly. 

The electrolytic refining of bullion has only been practi- 
cally carried out with the sulphate and nitrate baths, mainly 



ELECTROLYTIC REFINING OF DORE HULLION. 155 

the nitrate, which is in use in several large plants refining 
from perhaps 20,000 to 100,000 ounces per day. In either 
case the deposited silver comes down in a loose crystalline 
form. The older Moebius apparatus * is well described and 
illustrated in the patent specification and in several avail- 
able works. | The more recent Balbach apparatus, % im- 
proved by Mr. Wm. Thum, accomplishes the same result in a 
somewhat different manner. The following quotation and fig- 
ures are from Mr. Easterbrooks' paper, § read before the Ameri- 
can Electrochemical Society. 

"With electrolytic parting we have a choice of two dis- 
tinct systems of depositing silver on the cathode, one in a 
loose crystalline form at a relatively high current density, 
as in the Balbach and Moebius methods, the other with the 
aid of gelatine in an adherent form at a lower current density. 

"The electrolyte used is a copper-silver nitrate solution, 
although recently Betts |[ has proposed using a silver methyl- 
sulphate solution. 

"These methods all have in common the characteristic 
of parting and refining bullion free from gold and tellurium 
at one operation, the deposited silver being melted and poured 
into bars without any further refining, as in the sulphuric 
acid process. Silver placed in the tanks as anodes is not 
handled until taken out as refined silver, whereas in the acid 
method the silver either in solution or as cement must be trans- 
ferred several times with the aid of siphons, steam, etc., before 



* U. S. patent 310302 and 310533. Jan. 6, 1885. 

t Borchers, "Electric Smelting and Refining/' 2d Eng. Ed.; Watt and 
Philips '• Electroplating and Electro-refining," "Mineral Industry," 
Vol. VIII (1889), page 337. 

t U. S. patent 588524. 

I Trans. Am. Electrochemical Society, Vol. VIII (1905), page 131. 

II Electrochem. Industry, April, 1905. 



156 



LEAD REFINING BY ELECTROLYSIS. 



it is in a condition to be melted. For these reasons it is possible 
to operate an electrolytic parting plant with a higher degree 
of neatness and cleanliness (such as the value of the material 
treated requires) than is possible with acid parting. 

"A parting plant using the Balbach method is simple 
in construction and operation. Fig. 23 shows the cross- 
section of a tank. The cathode is made of one-half inch Acheson 
graphite slabs fitted to the bottom. Two silver contact-pieces 
rest respectively on the bullion to be parted and the graphite 
slabs. Bullion cast in thin square slabs is contained in a cloth 




Fig. 23. 

case which is supported on a wooden frame suspended over 
the tank. The gold slimes accumulate on the under side of 
the bullion, between it and the cathode, increasing the resist- 
ance as the operation continues. Each tank has a cathode 
surface of 8 square feet and a current density of 20 to 25 
amperes per square foot used.* The voltage averages 3.8 
per tank, and an average ampere efficiency of 93% was 
obtained on a continued run, while occasionally an efficiency 
of 98% was secured. The power required is 31.5 watt-hours 
per ounce of fine silver produced. 

"Most of the silver is deposited on the cathode surface 
directly under the anode, and the reduction of the distance 



* The U. S. Metals Refining Co. uses 50 amperes per square foot, 250 
amperes per cell. Engineering and Mining Journal, May 25, 1907, page 
1004. 



ELECTROLYTIC REFINING OF DORE BULLION. 



157 



betweeD anode and cathode is limited by the space necessary 
to roach in and remove it, which has to be done frequently 
on account of the silver bridging across to the cathode. This 
serves also to agitate the electrolyte. There is gassing in this 
tank and the consumption of nitric acid is much higher than 
in the Moebius method. 

"At 20 amperes per square foot about 32% of the daily 
output of each tank is held permanently in stock in electrolyte 




Fig. 24. 



and. contacts, which is less than is retained in the Moebius 
method. 

"In Fig. 24 is shown the cross-section of a Moebius 
tank. They are arranged in units of six placed end to end, 
each unit being provided with apparatus for raising the boxes 
containing the deposited silver together with the anodes and 
cathodes, and with arrangements for imparting a reciproca- 
ting motion to the wooden scrapers. There is no system of 



158 LEAD REFINING BY ELECTROLYSIS. 

circulating the electrolyte, but the scrapers moving back and 
forth agitate it. The anodes are contained in a cloth frame 
which holds the gold slimes, and the silver is brushed off from 
the silver cathodes by the wooden scrapers, and drops into 
a box with hinged bottom. It is removed by raising the boxes 
above the top of the tanks and emptying it into a tray placed 
beneath. This operation requires one-half hour per day per 
unit. Each tank has a cathode surface of about 16.5 square 
feet, and a current density of 20 to 25 amperes per square 
foot is used. The voltage between electrodes is 1.4 to 1.5 
and the power cost is 13.2 watt-hours per ounce of silver de- 
posited. An average ampere efficiency of 94% is obtained. 
At 20 amperes per square foot 41% of the daily output of 
each unit is permanently in stock in cathodes and electrolyte. 

' ' The necessity of cutting out of- service the units of a plant 
using the Moebius method to remove the silver, and the fre- 
quent siphoning off and replacing of portions of the electrolyte in 
each tank, in both the Balbach and Moebius methods, to main- 
tain it of fixed compositions, are objections overcome by de- 
positing the silver on the cathode in an adherent form. 

"This method permits of an arrangement of tanks and 
electrodes and a system of circulation of electrolyte similar 
to that used in the multiple system of copper refining. 

"The finely divided condition of the gold in the bullion, 
which in commercial work rarely contains more than 40 parts 
per thousand, requires the anodes to be inclosed in a cloth 
frame to keep the deposited silver free from gold, as the light, 
fine particles do not fall to the bottom of the tank with suffi- 
cient rapidity. A current density of 10 amperes per square 
foot is used, and the power cost is nearly identical with the 
Moebius method. Twenty-eight to 32% of the daily output 
is retained in cathodes and electrolyte." 

The last paragraphs refer to the refining of silver with 
the nitrate electrolyte, with the addition of gelatine to the 
solution, for the production of a solid cathode deposit. Mr. 



ELECTROLYTIC REFINING OF DORE BULLION. 159 

Easterbrooks exhibited some quite solid and very brittle ca- 
thode silver, with a nearly smooth surface. 

In the Philadelphia mint,* clore bullion containing 30% 
of gold is now refined clectrolytically with a solution contain- 
ing 3% of silver nitrate and 1J% of nitric acid, to which a 
little gelatine is added. Each cell is 40 ins. by 20 ins. and 
11 ins. deep, in which are hung 42 anodes 7 J ins. long, 
2V ins. wide, and | ins. thick, and 40 cathodes of the same 
width and length, rolled to 0.016 inch thickness. A current 
density of 7 amperes per square foot is used. From the 
above figures it is apparent that an electrode separation of 
3 inches or more must be used, which is more than would 
be necessary if the silver actually comes clown solid. The 
photograph showed the character of the deposit, which prob- 
ably consists of a large number of roundish masses of silver 
lightly fastened together, but with sufficient tenacity to keep 
from dropping into the cells to any serious extent. 

The Moebius and Nebel process, using a traveling silver 
belt to collect the silver, is variously described.! The article 
by Mr. M. W. lies J in "The Mineral Industry" gives a rather 
full description of the plant with observations on the amount 
of nitric acid used; construction of the gold room; inventory 
of gold and silver; action of nitric acid on the silver belts; 
testing of the solution; silver vs. platinum contact-points 
for the anodes, and costs, as follows: 



* Annual Report of the Director of the U. S. Mint, 1905, abstracted in 
"Electrochemical and Metallurgical Industry," 1906. Vol. IV, page 306. 

t English patent 469 of 1895, January 8th. U. S. A. patents 532209 
January 8, 1895; 592097, October 26,1897; "Electroplating and Electro- 
refining," Watt and Philip, page 576; Borcher's "Electric Smelting and 
Refining," 2d Eng. Ed., page 323. 

t "The Mineral Industry," page 337. Vol. VIII. 



160 LEAD REFINING BY ELECTROLYSIS. 

TABLE 62. 

Supplies. Per Month. 

Oil $56 . 20 

Nitric acid, 1698 lbs. at 7.5 cents 127.35 

Waste, 113 lbs. at 9 . 5 cents 10 . 73 

Coal, 25.47 tons at $2.25 57.31 

Coke, 1543 lbs. at $9 . 50 7 . 32 

Cupels for melting silver 6 . 75 

Crucibles, 2 No. 40 at $2 4 . 00 

Sundry supplies 21 . 12 

$290.78 

Labor. Per Month. 

Assistant superintendent $160 . 00 

Five men 31 days 379 . 75 

Superintendent half time 200 . 00 

$739 . 75 
Interest $200,000 at 10% $1,666.67 

TABLE 63. 

Cost per Ounce of Bullion. 

Supplies 0427 cents. 

Labor 1087 ' ' 

Interest 2450 " 

Total 3964 cents. 

Royalty 1000 ' ' 

.4964 cents. 

The article concludes with a statement that the cost could 
be considerably reduced. The rate of interest charged was 
particularly high. 

Through the kindness of the Compania Minera, Fundidora 
y Afinadora, Monterey, Monterey, Mexico, Mr. A. K. Brewer, 
Superintendent, I am able to give a photograph of their parting 
plant, Plate 3, and accurate information regarding it as follows: 

Capacity of the plant is 1000 kilos = 32,150 ounces per 
twenty-four hours. The dore runs from 985 to 992 parts 
per thousand in silver and gold, the gold making up from 
2 to 60 parts of the total. The 48 tanks take 250 amperes, 




H 5 



o < 

H |8 



ELECTROLYTIC REFINING OF DORK BULLION. 163 

at 2 volts per tank, equal to 24 K.W. for the whole plant. 
Five horse-power is used in addition to drive the belts, revol- 

ing brushes, and solution pump. The circulation of the 
electrolyte^ is perfect and flows from an upper storage-tank 
through the 4 cells and into a tank under the floor, whence it 
is raised by the pump to the upper tank. To maintain the 
solution a few barrels of it are occasionally removed, and 
added to the ore-beds, so that the values gc through the 
smelter. 

The electrolyte contains 20-50 grams silver, 10 to 20 
grams copper, 2.5 to 15 grams lead, and 2.5 to 10 grams free 
nitric acid per litre. Nitric acid is added from time to time 
to the solution in the lower storage-tank to maintain the 
electrolyte at working strength. 

Each tank takes 22 anodes 3 ins. by 12 ins. by i to J ins. 
thick, which weigh 0.5 to 2 kilos apiece, so that the amount 
of silver in the tanks is probably about | to 1 day's output. 
There is no anode scrap, the anodes being totally dissolved, 
except the gold. The consumption of nitric acid is about 
40 lbs. for 32,000 ounces dore. Men required are three day- 
times and two at night, including foreman and melter. 

It is possible to form a close estimate of the cost of parting 
with this apparatus, on the above results. 

TABLE 64. 

Per Oz. 

Power at S60 per E.H.P. vear would be * 0190 cents. 

Labor at S3 average * 0470 ' ' 

Nitric acid at 5 cents per lb 0060 ' ' 

Interest on dore in tanks at 85 cents per oz 0142 ' ' 

Interest on other gold and silver 0284 ' ' 

Interest on plant, including solution 0090 ' ' 

Fuel and materials for melting 0100 " 

Superintendence 0120 ' ' 

. 1456 cents. 
* Assumed. 



164 



LEAD REFINING BY ELECTROLYSIS. 



The costs can not be directly compared with those given 
below for other methods because of larger scale of opera- 
tions. Refining 20,000 ozs. per day, the superintendence 
and labor items would be quite a little higher per ounce, say 
.019 cents for superintendence, and .063 for labor. Allow- 
ing for cost of new belts occasionally, the total cost on a 
scale of 20,000 ounces per day would approximate to .16 to 
.17 cents per ounce. 

The following description and drawing (Fig. 25) of the 
Moebius and Nebel apparatus are taken from their U. S. 
patent : 

Referring now to Fig. 25, the letter A designates the elec- 




Fig. 25. 



trolytic tank, made by preference of a solid block of wood 
dug out and suitably lined. 

BB' are rolls adjustably mounted in brackets placed on 
the tank; CC, an endless silver cathode-belt passing over 
the rolls BB'. 

W are the shafts of the rolls BB', mounted in brackets 
dd' and adjusted by screw-bolts gg', so as to impart to the 
belt the proper tension. 

DD' are rolls to keep the part C of the belt immersed in 
the bath, the roll D being formed with teeth, as shown, so 
as not too much to press down the silver precipitated thereon. 
The roll D' may have a plain cylindrical surface. 



ELECTROLYTIC REFINING OF DORE BULLION. 165 

Slow motion in the direction of the arrows is imparted to 
the belt CC by any suitable means, such as the sprocket- 
wheels Ww and chain m, operated by a belt-pulley mounted 
on the shaft s of the small sprocket-wheel w. 

T is a circular brush held against the belt while passing 
over the roll B and by a weighted arm pp', mounted loosely 
on the shaft s, the brush being actuated from the shaft s by 
suitable gear, so as to brush the silver from the belt into the 
receptacle R. . 

U is an oil-tank, within which are mounted two rolls u 
and r, both of them a little longer than the width of the belt. 
As shown, the oil-tank is suspended from the bracket df in 
such a manner that both rolls u and r are continuously pressed 
against the belt. The roll u is rotated by contact with the 
lower part C of the silver-belt and oils the surface of the 
same, upon which the silver is afterward deposited when in 
the position C. The roll r is normally held by a pawl t and 
serves to remove or scrape off any surplus of oil. By raising 
the pawl t the roll r may be revolved, so as to remove any 
matter that may have been accumulated thereon. The 
roll r is, by preference, made of material such as lampwick 
properly secured to the shaft in the usual manner. Any other 
suitable oiling apparatus may be used. 

The letter E designates one of the anode-cells, the anode 
being connected to the conductor K, while the belt is con- 
nected to the conductor L by a brush F. 

A great many experiments have been made in my labor- 
atory with the aim of finding a process by which silver could 
be refined in the same manner that copper and lead are, with- 
out the use of any special arrangement to collect cement silver, 
but to deposit solid silver on the cathodes at once. 



166 LEAD REFINING BY ELECTROLYSIS. 

A number of other objects were in view at the same time. 
One was to use a solution which would take any bismuth in 
the anodes into solution. Another was to use a more highly- 
conducting solution and use higher current density, thus cut- 
ting down power and interest. 

The best deposits were got with a solution of silver methyl- 
sulphate. The deposit of silver was adherent and dense, but 
not entirely solid. 

The silver methyl-sulphate electrolyte in distinction from 
the nitrate electrolyte, can be made strongly acid, and hence 
highly conducting, a very important advantage in silver re- 
fining, as it permits higher current densities. The other elec- 
trolytes tried, of silver dithionate and fluoborate, though 
strongly acid and excellent conductors, would not dissolve 
bismuth in quantity and gave somewhat inferior results in 
other respects. 

Experiments were also made with amyl-sulphate solutions 
strongly acid from amyl-sulphuric acid, with and without 
the addition of gum arabic, etc., and it appeared that there 
was a point to be reached in respect to strength of solution 
and percentage of gum arabic, etc., where the deposit was 
neither bright and loose, nor dark and soft, but smooth and 
fairly solid. 

The deposition of entirely solid silver requires a delicate 
balance of conditions, and some unexplained phenomena 
must have presented themselves to experimenters. One 
curious fact is that a silver electrolyte has to be in use for 
a considerable number of hours before it gets into good work- 
ing order and the results at the cathode strongly resemble 
those obtained in starting up with a new lead solution, when 
traces of arsenic and antimonv come down with the lead and 



ELECTROLYTIC REFINING OF DORK BULLION. 167 

make it impossible to get a solid deposit. I think it probable 
that tlu^ same thing occurs in the case of silver — that the 
preparations of silver carbonate, silver nitrate, and silver 
sulphate, etc., used in making up solutions, contain traces 
of other elements which deposit with the silver and spoil it 
mechanically. Possibly a trace of platinum is what does it r 
or perhaps a modification of silver itself. It is known that 
sometimes more silver deposits than is demanded by theory, 
and it has been suggested that this is due to the deposition 
of colloidal particles of silver. In support of the above ideas, 
at one time I prepared a solution for depositing silver, by 
electrolysis of the solution with a silver anode in a diaphragm- 
cell. The resulting solution was one of silver methyl-sul- 
phate, and gave a beautiful bluish, smooth, solid deposit of 
silver, not inferior in structure to electrolytic copper. 

The use of a higher anode current density, that is, above 
say 20 or 30 amperes per square foot, is undesirable with the 
methyl-sulphate solution. On one occasion a methyl-sulphate 
solution that was yielding a dense deposit of silver gave a 
very poor deposit soon after the substitution of a smaller 
and purer anode, the current and cathode area remaining the 
same. 

The use of perchlorate of silver, which has been used by 
Carhart, Willard, and Henderson * in the silver coulomb- 
meter with much better results than were formerly obtained 
with silver nitrate, is analogous to the use of methyl-sulphuric 
acid, as it is also a strong acid that can be used in large excess 
above that required to dissolve the silver. 

Methyl-sulphuric acid is prepared by mixing together 

* Am. Chem. Soc, Vol. IX, page 395. 



168 



LEAD REFINING BY ELECTYOLYSIS. 



methyl alcohol and sulphuric acid. The mixture heats up, 
and the reaction only takes a short time. Previous results 
that indicated a period of eight to ten hours' reaction at 100° C. 
and statements in text-books to the same effect are wrong, 
and it is doubtful if the reaction takes more than time 
enough for mixing. 

I had experiments made in my laboratory with various 
mixtures of 96% sulphuric acid and 88% methyl-alcohol, 
with different heat treatment. The best results were got by 
simply adding the alcohol to the acid, mixing well, allowing it 
to stand five minutes, and pouring into cold water (pouring 
on ice would be better in practice). 

The results in that way were as follows: 

TABLE 65. 



No. 



H:S04 



Wood Alcohol. 



1 


20 cc.=35.8 gr. H,S0 4 


15 cc. = 10.9 gr. CH 4 


2 


20 cc. = 35.8 gr. H,S0 4 


12 cc.= 8.8 gr. CH 4 


3 


20 cc.=35.8 gr. HoS0 4 


10 cc.= 7.3 gr. CH 4 


4 


20 cc. = 35.8 gr. H.,S0 4 


8 cc.= 5.85 gr. CH 4 


5 


20 cc.=35.8 gr. H^0 4 


6 cc.= 4.4 gr. CH 4 



No. 



H2SO4 Utilized. 



Alcohol Utilized. 



48% of total 
42% " 
38% " " 
33%- " 
30%^ " 



51% of total 
56%- " 
61%" - 
66%- - 
80%- - 



With C.P. methyl-alcohol, specific gravity .817 = 92%, the 
result was as follows: 







TABLE 66. 


H2SO4 

Utilized. 


Alcohol 
Utilized. 


20 cc. = 35.8 gr. 


H 2 S0 4 


12cc. = 8.9 gr. CH 4 


47% total 


61% total 


20 cc.=35.8 gr. 


H 2 S0 4 


10cc.=7.44 gr. CH 4 


42% - 


65% - 


20 cc. = 35.8 gr. 


H 2 S0 4 


8 cc.=5.95 gr. CH 4 


37% - 


72% - 



ELECTROLYTIC REFINING OF DORE BULLION. 169 

The formation of water prevents complete reaction. The 
materials used in the experiments already contained water. 
With anhydrous materials the results must be better. Bet- 
ter results still are got with fuming H 2 S0 4 , which is now pro- 
curable at about 1.3 cents per lb. for acid containing 30% 
of S0 3 . Experiment: 155 grams fuming H2SO4 containing 
}i) ( c S0 3 , 20 cc. concentrated H 2 S0 4 , and 75 cc. wood alcohol 
. added together in small portions, first one and then 
the other, starting with alcohol and ordinary H2SO4, gave a 
yield of 53% on the acid and 67% on the alcohol. For com- 
parison with the above results, the proportions of S0 3 and 
alcohol in this experiment are the same as with 20 cc. 96% 
H0SO4 and 13.5 cc. alcohol, when the yield on acid is about 
46% and on alcohol, say 51%. 

The course of the reaction was traced by titrating a 
sample with ammonia and cochineal. Acid disappears in the 
reaction, as one molecule of dibasic acid produces one molecule 
of a monobasic acicl, and the amount shown by titration to have 
disappeared multiplied by two gives the amount of acid utilized, 
from which can be calculated the amount of alcohol combined. 

In practice the product is poured on ice and the liquid 
treated with lead carbonate (though lime or baryta would also 
do) in amount sufficient to remove all H2SO4. The filtrate 
from the lead or calcium or barium sulphate is then treated 
with silver carbonate (from silver sulphate and soda) when 
the solution is ready for use if of the right strength, namely, 
about 15% CH3SO/ and 4-6% Ag. 

Probably ethyl alcohol can be used equally as well, but con- 
sidering the relative molecular weights ethyl alcohol would 
have to be 1.425 times as cheap as wood alcohol, to compete. 

A current density of 20 to 30 amperes per square foot is 



170 LEAD REFINING BY ELECTROLYSIS. 

permissible, and the solution, with agitation, may be reduced 
to 1.5 grams of silver per 100 cc. before it is necessary to 
strengthen it up again. The addition of gelatine or other 
materials is not recommended at present, as the} T . are hard 
to control in their action and the deposit is as satisfactory 
without. The anodes should be wrapped in cloth. Silver- 
plated and slightly greased graphite cathodes may be used 
to advantage, to which the silver adheres though not very 
securely. After one day's refining the cathodes are removed 
and the silver split off and the cathodes returned to the bath. 
As some silver is likely to be knocked off in the cells, the use 
of storage-battery glass cells is convenient. These can be 
handled and cleaned easily, and will take a fairly large cur- 
rent. A cell about 12" square and 15" deep can easily take 
110 amperes, and perhaps as high as 200, while a stoneware 
Balbach cell about 4 feet long, 1 foot deep, and 2 feet wide, 
is only good for about 100, perhaps 200 amperes, and takes 
up eight times the space. 

The cost of refining by the various electrolytic methods can 
be estimated as follows, from various data. In all cases the in- 
terest on the original cost of plant is taken at 109c and on metal 
on hand at 6%. It is evident that the cost of melting dore bul- 
lion and refined silver will be practically the same in all cases. 

Comparative cost, refining 20,000 ozs. per day, Table 67. 

TABLE 67 

Cents per Oz. 

Moebius. Balbach. Betts. 

Interest on plant, including solution 0090 . 0088 . 0042 

Power at S60 per E.H.P. year 0146 .0369 . 0049 

Interest on dore, in cells at $0.85 oz - . 0142 . 0142 . 01^2 

Interest on other gold and silver in stock ■ 0284 . 0284 . 0284 

Labor and superintendence 0850 . 0850 .0850 

Chemicals 0100 .0150 .0100 

Fuel and material for melting 0100 . 0100 . 0065 

.1712 .1983 .1532 



ELECTROLYTIC REFINING OF DORK BILLION. 171 

The above figures can be expected to be fairly close, but 
the fact that the Balbach method, as modified by Mr. Wm. 
Thum,* has been recently introduced *in new plants, speaks 
against the above figures. It is difficult to see wherein the 
new process has the advantage over the Moebius, unless in 
the matter of labor cost or possibly interest on dore in the 
rolls. It seems, however, probable that the above figures for 
the Balbach process are a little too high. One manager 
remarked to me that he thought the cost of operating the 
Moebius and Balbach process about the same, with the 
advantage of simplicity in favor of the latter. 

When it comes, however, to refining dore bullion con- 
taining important quantities of base metal, as copper, lead, or 
bismuth, the results are somewhat different, and can be best 
expressed by a formula of the form 

C = A+xC + yB+zP, 

in which A is the cost of melting and refining an ounce of 
dore free from base metals, and C, B, and P are the costs of 
recovering from the electrolyte, as marketable metal, one 
ounce each of copper; bismuth, and lead respectively, and 
adding the equivalent of silver to the solution, while z, y, 
and z are the respective proportions present. 

In the Moebius and Balbach processes the cost of recover- 
ing bismuth per Troy ounce from the anode slime would approx- 
mate as follows: 

For washing gold with soda to form the soluble variety of bis- 
muth hydrate or carbonate, and dissolving in cold nitric acid 



*U. S. Patent. 



172 LEAD REFINING BY ELECTROLYSIS. 

heating the solution to precipitate basic nitrate, about 0.13 

cent 0.13 cent. 

Converting basic nitrate to metal by smelting with charcoal, 

about 0.13 cent 0.05 " 

Silver carbonate to make up for weakening of electrolyte 0.48 " 

Ttoal 0.64 cent. 

For copper, copper nitrate can be crystallized out and 
this could be electrolyzed in a dilute solution for copper and 
nitric acid, and the nitric acid returned to the bath, though 
this is not probably actually done. 

Estimated cost per Troy ounce copper in bullion 3 cent. 

If lead nitrate crystallizes with the copper nitrate, evi- 
dently the two may be dissolved together and the copper de- 
posited out with platinum or carbon anode as above, while 
the residual lead nitrate can be crystallized from the mother 
liquor, lead peroxide being also produced, however. 

Estimated cost of evaporating lead nitrate per Troy ounce and 

corresponding loss of nitric acid 3 cent. 

If the dore should contain then 10% lead, 10% bismuth, 
and 10% copper, the cost per ounce ought to approximate 
to the result given by the formula above. 

C=.18+Jq.64+yq.3+Jq.3 = .304 cent. 

With the Betts parting process the values would be some- 
what different, and considerably lower, for (l) there is no 
appreciable loss of the acid making the basis of the electrolyte, 
(2) no separate operation for removing bismuth from the 
gold slime, and (3) the working up of the copper-silver pre- 



ELECTROLYTIC REFINING OF DORE BULLION. 173 

cipitatc thrown out by metallic bismuth and the bismuth 
and copper thrown out by metallic lead, by treatment with 
ferric sulphate, hot sulphuric acid, etc., is simpler and direct. 
I should estimate the values for C, B, and P at .5 cent, 
.2 cent, and .1 cent, respectively. If these values are realized, 
the cost for the same dore bullion would be 

Cost = .17+y^.5+jQ.2 + — 1 = .215 cent. 

These results are not intended to be entirely accurate, 
and of course they can not be. 



CHAPTER V. 

THE MANUFACTURE OF HYDROFLUORIC AND FLUOSILICIC 

ACIDS. 

*"No very useful literature on this subject exists to the 
best of my knowledge. Most chemists regard it as an ex- 
tremely dangerous substance, and have presumably left it 
alone as much as possible. Yet hydrofluoric acid and fluo- 
rides have an extending use for numerous purposes. Its 
preparation is easy and safe, if proper precautions are taken. 

" Samples of fluorspar may be tested by mixing say 50 
grams with various proportions of 66° sulphuric acid in small 
sheet-iron pans and distilling under the hood. For prepara- 
tion in small quantities for the laboratory, apparatus as shown 
in Fig. 26 gives good results if used out of doors. The retort 
is an ordinary cast-iron pot, perhaps one foot in diameter 
and 6 inches deep. The cover is made by filling with sand 
to near the top, leveling it off and pouring in about \ inch 
of lead. The lead pipe is separate from the cover, and passes 
over to a lead hydrofluoric-acid bottle containing water. The 
water must not come as high as the end of the lead pipe. 

" During distillation the bottle is sprayed with water from 
a hose to keep it cool. A charge of about 2 kg. of fluorspar 
and 2.5 kg. H 2 S0 4 66°, is stirred up in the pot. The 

* By permission of the Engineering and Mining Journal, April 20, 
1907. 

174 



HYDROFLUORIC AND FLUOSILICIC ACIDS. 



175 



fluorspar, for the most part, dissolves immediately on stir- 
ring in the sulphuric acid, without evolution of much fume, 
until heat is applied. The cover is put on and dry cement 
put over the joints as a lute, cement being suitable for this 
purpose. 

"The heating should be moderate at first to prevent too 
much frothing in the pot. Distillation takes two or three 




Fig. 26. 



hours, and the end can be told by feeling of the lead pipe near 
the bottle, which is hot as long as acid is coming over. Very 
little loss is experienced and a yield of 80% or thereabout, 
is obtained. 

" Operation on larger scale. — On a large scale, the applica- 
tion of the same principles is successful. The general arrange- 
ment is shown in Fig. 27, for which a few explanations are 
necessary. The pot may be cast about 8 ft. in diameter, 3 ft. 
deep at the center, and 1 in. thick, with a slightly curving 
bottom to prevent cracking. For the pot a cast-iron cover 
1 in. thick is used, dipping into the annular trough around 
the pot, which contains strong sulphuric acid as a seal. All 
the other seals are made in the same way, but water may be 
used for the joints on the condensers where the temperature 



176 



LEAD REFINING BY ELECTROLYSIS. 



is not so high. Lead retorts and lead covers for the retorts 
are useless. 

"The condensers consist of a series of two or three lead 
boxes of about 1 cu.m. capacity, entirely submerged in a 
water-tank and partially filled with water or dilute HF. Con- 
densers should be made of heavy lead, supported by wooden 
pieces to which the lead is attached by means of lead straps 
burned on. The lead delivery-pipes may be about 5 in. in 
diameter. The condensers have an overflow so that the acid 



S 



M 



rf 



\3- 



T= 




Fig. 27. 



never can rise to the end of the delivery-pipe. If this hap- 
pened, a partial vacuum might result, and draw water back 
into the pot, where it would probably cause an explosion. 

"The charge may consist of 1000 lbs. of ground fluorspar 
and 1000 to 1200 lbs. 66° sulphuric acid. SiF 4 comes off 
first and deposits silica on the water in the first condenser, 
stopping absorption somewhat, so that it is necessary to stir 
the water in the first condenser until most of the SiF 4 has 
come over. The pot may be charged in the morning and 
distillation finished by night. Coal is used for fuel, burned 
on a grate ot about 3 square feet. The residue in the pot is 
comparatively hard, and, after cooling, is dug out with pick 
and shovel. The yield of acid calculated on the sulphuric 
acid used is approximately 80 to 90%. 

. "The cost of manufacture is not great, the principal 
items being the raw materials necessary. To produce 1 lb. 



HYDROFLUORIC AND FLUOSILICIC ACIDS 177 

anhydrous I IF, about 2 J lbs. of fluorspar and 3 lbs. sulphuric 
acid are necessary. Fluorspar and sulphuric acid are worth 
about $10 to $15 a ton, making a cost for raw materials, 
exclusive of coal, of approximately 2f to 4 J cents per pound 
anhydrous HF. 

"Method of analysis. — The sample of acid is mixed with 
several times its bulk of nearly saturated and neutral potas- 
sium nitrate solution. This causes a precipitation of potas- 
sium fluosilicate in the solution: Phenolphthalein is used as 
indicator, and the solution titrated with caustic soda in the 
cold. This gives the total of the HF and H 2 SiF 6 present. 
The sample is then heated to boiling, when it will be found 
that considerable more caustic soda may be run in to get 
another end point. In the first titration, the HF present and 
the HN0 3 liberated by the reaction of potassium nitrate and 
fluosilicic acid are neutralized by the alkali. When titrated 
hot, the precipitated K 2 SiF 6 is decomposed by the alkali. 
The following is the equation involved: 

K 2 SiF 6 + 4NaOH = 2KF + 4NaF + Si0 2 + 2H 2 0. 

"The rule for calculating is, 1 gr. NaOH used in the second 
titration = 0.9 gr. H 2 SiF 6 in the sample. For HF present 
divide the number of cubic centimeters of NaOH used in the 
second titration by 2, and subtract the result from cubic centi- 
meters used in the first titration. The remainder shows the 
HF, 1 gr. of NaOH equalling 0.5 gr. HF. 

" Hydrofluoric acid has been shipped in beer-barrels with 
rosin lining, which are entirely successful, and last for some 
time and for long shipments; also in rectangular lead carboys. 
Its storage in lead is not very satisfactory on account of the 



178 



LEAD REFINING BY ELECTROLYSIS. 



corrosion of the lead. Probably the presence of sulphuric 
and fluosilicic acids has some effect in the corrosion. 

"I am indebted to Dr. William Valentine for some of my 
data." 

The conversion of hydrofluoric acid to fluosilicic acid can 
be accomplished in a lead-lined tank as shown in Fig. 28. 

The tank may be made about 5 or 6 feet square and is 
one-third filled with clean sand or broken quartz. The method 
of operation is based on the discovery that while cold hydro- 




fluoric acid will pass through sand and be only partly con- 
verted to H 2 SiF 6 , if the acid is hot, the reaction will easily 
maintain the heat and pure H 2 SiF 6 will run through. Accord- 
ingly on the start the tank is filled with water and steam blown 
in to heat it to boiling. When the water running through 
begins to get hot, it is allowed to drain off, and 30-35% acid 
added. The tank is kept covered by boards, but acid would 
boil off in large quantities, except for the addition of cold 
water in sufficient amount to prevent this. As the acid runs 
out of the tank (one square foot of sand at Trail used to let 



HYDROFLUORIC AND FLUOCILICIC ACID. 179 

acid through at about the rate of one barrel in twenty-four 
hours) more is added, with enough cold water to prevent boil- 
ing off of acid. As long as the supply of acid is maintained 
the tank will not cool off, and the acid running through has 
only to be diluted and have white lead added. 

The tank should be elevated so that the products can run 
off into other tanks. At Trail the acid was hoisted to the 
tank in barrels, the bung knocked in, and the acid poured 
into the tank. This was a very disagreeable job. A lead- 
lined montejus, if a supply of acid under pressure is available, 
would be much better to work with. When convenient the 
hydrofluoric acid is made on the hillside above the works, 
so that it may be entirely managed by gravity. 



CHAPTER VI. 

CHOICE OF CONSTANTS. 

This chapter will be a study of the relative advantages 
of various rates of working, arrangement of plant, methods 
of slime treatment, etc. 

Probably the chief point to be decided is the current 
density to be used in depositing the lead. The problem can 
be looked at on many sides, but most of these can be elimi- 
nated at once as having no real influence on the result. 

There is the choice to be made between the series and 
multiple arrangements of electrodes. The important advan- 
tages of the two are probably as follows: 

Series system. — Power cost about 40%-50% less, or a 
saving of about 34 K.W. hours, worth about 28 cents per 
ton. 

No starting sheets required, or a saving of about 15-20 
cents per ton over lead cathodes, and much less over lead- 
plated steel cathodes. 

Smaller construction cost for plant, excluding power plant, 
of "about S50 per ton per day, or at 10% per annum for in- 
terest, 4 cents per ton. 

Total of advantages, about 50 cents per ton. 

To offset this, the multiple process will require only about 
half as many anodes cast and charged, and will produce less 
anode scrap, an advantage of 10-15 cents probably. 

180 



CHOICE OF CONSTANTS. 181 

No necessity of separating anode scrap and slime from 
cathode lead, an advantage that might easily be 20 cents 
per ton and probably more, while producing better refined 
lead, too. Less interest charge on anodes, which might easily 
be about 5 cents per ton. Total of advantages, 35-40 cents 
per ton, or more. 

There are probably other disadvantages connected with the 
series system that are only familiar to those who have had 
experience with it. 

The character of the bullion would have to be carefully con- 
sidered in this connection. The series process would succeed 
best with lead bullion giving very little slime, such as Missouri 
lead or relatively impure bullion containing \\% of antimony 
and arsenic or more. With this latter kind of lead the slime re- 
mains closely adherent, and probably the entire anode could be 
dissolved through and the process stopped when the cathode 
lead on the other side was being first attacked. The slime 
would remain as a soft, porous slab separate from the cathode 
lead. With the average grades of bullion containing little 
arsenic and 0.5-1% antimony, the slime is so voluminous 
and soft that it would be apt to slip off the anode and fill 
most of the space between the electrodes, which condition 
could perhaps be remedied by making the tanks nearly twice 
as deep as the electrodes, with boards placed across the tank 
every foot or so to prevent the current passing through the 
bottom part to a large extent. 

The published information on the series process as applied 
to copper is not entirely applicable to lead. One of the great 
objections to the series system in copper refining is the cost 
of making smooth and uniform electrodes. This would be 
much easier with lead, on account of the greater facility of 



182 



LEAD REFINING BY ELECTROLYSIS. 



melting and rolling it. The separation of anode scrap from 
deposited metal, said to cost 60 cents per ton with copper, 
would not be nearly so large an item with lead. 

With little probability of doing much better with the series 
system, and some of doing worse, there is little chance of its 
being attempted, except for small plants in which it is never 
convenient to generate large currents of high amperage and 
low voltage, or where power is very expensive. 

Even for small plants another system proposed,* which 
combines to some extent the advantages of the two systems, 




Fig. 28a. 

is apt to be better than the series system. The principle of 
this arrangement can be easily noted from Fig. 28a. 

With certain improvements in the multiple process that 
seems feasible, and are noted elsewhere, the multiple process 
would have a decided advantage over the series process. 

We have these points to consider in choosing current 
density: 

(1) Purity of the lead. — Obviously the current could be so 
high as to dissolve impurities which would deposit on the 



* U. S. Patent 789353. May 9, 1905. 



CHOICE OF CONSTANTS. 183 

cathodes. This is not a factor having any important influ- 
ence in choosing the current density, as it has been amply 
demonstrated that pure lead can be produced over any range 
of current density that is permissible from other considera- 
tions. 

(2) Cost of glue used in the solution. — It seems probable that 
the consumption of glue increases with increase of current 
to some extent. As the amount of glue used per ton of 
lead produced is only about one-half to three-quarters of a 
pound, it will be seen that a small increase or decrease of 
this amount is too small a factor to be considered. 

(3) Low current density means larger tank room and 
consequently somewhat greater cost of building. Each am- 
pere per square foot below 12 amperes, would make an extra 
cost of building of approximately $20 per ton refined per 
day, while increasing the current up to say 20 amperes would 
save approximately $100 per ton per day on this score. 
Capitalized at 10%, the total difference between 10 and 20 
amperes is only about 3 cents per ton. In some circumstances 
the value of land will enter into the question, but in that 
event it may be better to economize in space by leaving smaller 
passages between the tanks, and making the tanks somewhat 
deeper. At the plant of Locke, Blackett & Co., Ltd., New- 
castle-on-Tyne, this method was adopted, and the current 
is 12 amperes per square foot, with a space between the rows 
of tanks and around the sides of about 20 inches. This is 
hardly as convenient, but it makes little if any difference in 
the cost per ton refined. 

(4) Interest on metal tied up. — We can calculate the thick- 
ness of metal dissolved per week at various current densities, 
at 95% efficiency, as follows, Table 68: 



184 



LEAD REFINING BY ELECTROLYSIS. 





TABLE 


68. 








Current Density Amperes 
Square Foot. 


per 


Inches of Anode Dissolved 
Week, on Each Side. 


per 


10 








.243 




12.5 








.305 




15 








.365 




17.5 








.425 




20 








.486 





The amount of metal tied up may be varied by varying 
the thickness of the anodes, but it is, of course, uneconomical 
to cast them very thin on account of the extra cost of casting, 
placing in tanks, cleaning, etc. The best thickness of anode 
is really dependent mainly on the thickness of the cathodes 
it is possible to make. Each set of anodes should be made 
so as to give either one or two sets of cathodes deposited as 
thick as practicable. Present experience indicates that cathodes 
with about 35 lbs. deposited per square foot are as heavy as 
it is desirable to make them. For an anode yielding two sets 
of cathodes, and allowing 15% for scrap to be remelted and 
slime, makes a 500-lb. anode with the usual size, 2 feet wide 
and 3 feet deep. With various current densities, bullion 
valued at $175 per ton, an average of five-sixths of the total 
value being in tanks, and allowing one day's supply unmelted 
cathodes and one in stock, and half day for melting each, the 
results are as follows: 



TABLE 69. 



Current Density. 


Ins. Dissolved per 
Week, Both Sides. 


Value of Metal on 

Hand per Ton 
Refined per Day. 


Int. Charges per Ton 
Refined at 6^- 


10 

12.5 
15 

17.5 
20 


.486 

.608 

.73 

.85 

.974 


$2570 

2160 
1885 
1690 
1550 


SO. 423 
0.355 
0.310 
0.278 
0.255 



CHOICE OF CONSTANTS. 185 

The interest charge with the low current density is quite 
considerable, and could be reduced to perhaps two-thirds 
that amount by casting anodes one-half as thick; but the 
extra c^st of casting and handling so many more pieces would 
leave little or no net saving. 

(5) Depreciation of tanks. — With the wooden tanks hitherto 
used, the life of which may be taken at four years, and cost- 
ing $40 each, evidently higher current density means the 
maintenance of fewer tanks in a direct ratio, about as follows, 
allowing for a certain amount of repairs: 

TABLE 70. 

Current Density. ^^^tetJt^ *" T ° n 
10 $0,143 

12.5 0.114 

15 0.095 

17.5 0.081 

20 0.071 

(6) At Trail the current density is about 16 amperes per 
square foot, and the total loss of acid is stated to be 10 lbs. 
of anhydrous H 2 SiF 6 per ton,* the solution containing 6-7 gr. 
lead and 12-13 gr. SiF 6 per 100 cc. At Newcastle-on- 
Tyne current density 11 amperes, the loss at one time was 
determined as 6 lbs. per ton of 2240 lbs., the solution con- 
taining 6 gr. lead and 15 gr. SiF 6 per 100 cc. 

The acid loss at Trail in 1902 and 1903 was as follows: 

TABLE 71. 

August and September 16, 1902 13.8 lbs. SiF 6 per ton lead. 

September 16-October 6, 1902 7.7 " SiF 6 " " " 

January 22-February 13, 1903 6.3 " SiF 6 " " " 

* Communicated by Mr. A. J. McNab, Trail, B. C. See Appendix. 



186 LEAD REFINING BY ELECTROLYSIS. 

In the last two determinations the current density was 
about 12 and 10 amperes per square foot respectively, with 
solutions containing 7.5 gr. and 8.5 gr. SiF 6 per 100 cc. 
respectively. These figures are still too high for good work, 
as the arrangements for saving leaks and wash-waters were 
crude. This should be noted especially for the first 
period, as operations had not been at all systematized 
then. 

The use of a high current density would tend to diminish 
acid loss from leaks, but there need be no loss from leaks any- 
way with good tanks and proper supports. The only way in 
which high current density could increase acid loss would be 
by depositing silica in the slime faster than the free HF in 
solution could dissolve it, but that means only a loss of the 
relatively valueless silica, which can be cured by dissolving 
fresh silica in the solution, or by stirring the slime up well 
with the electrolyte to secure a recombination of silica and HF. 
The latter simple and practicable procedure has not yet been 
introduced in practice, as far as I know. 

Lacking determinations of acid loss at varying current 
densities, and in view of the facts we have which do not make 
it seem probable that moderately higher current densities 
w T oulcl increase the acid loss, it would not be safe to speculate 
much on the effect of varying current density. 

(7) Interest on copper conductors. — This is not a variable 
in respect to current density to any extent, and need not be 
considered in the present inquiry. 

(8) Interest on tanks and electrolyte is a small item of a 
few cents only and not worth considering in this connection. 
The difference in the solidity of the lead and labor cost for 
keeping tanks in good working order is not considered to 



CHOICE OF CONSTANTS. 187 

vary appreciably with variation in current density, within 
the limits considered here. 

(9) Power. — The power cost per ton varies in nearly direct 
proportion to the current density, and also of course with 
the cost of electrical power, and this latter may be taken at 
150 per E.H.P. year, which seems a high enough average. It 
is now possible, by the use of gas-engines, water-power, or 
cheap coal, to generally reach or surpass this figure. It is also 
possible to secure wide variation in power cost, by varying 
the strength of the electrolyte. Inasmuch as even with a 
low current density of say 10 amperes per square foot, it is 
economy to use a rather strong solution containing about 
1G gr. SiF 6 " per 100 cc. (except where acid is unduly expen- 
sive), my figures are based on a solution of 7-8 gr. Pb" 
and 16-17 gr. SiF 6 " for various current densities, and also 
for comparison, with a solution containing 10 gr. lead and 
20 gr. SiF 6 . For conductivity determinations, see Tables 18 
and 19 and Figs. 2, 3, and 4. 

The temperature effect, although the conductivity varies 
quite a little with change of temperature, is not of much prac- 
tical importance. At Trail at one time the electrolyte was 
heated as high as 50°, by a steam coil in the circulation-tank, 
but the practice was found unsatisfactory in several ways, 
while the gain in conductivity was not large. 

The effect of temperature up to 30° C. is illustrated in Figs. 
3 and 4. The resistance was not measured at higher tem- 
peratures, but can be safely calculated to about 45° by extra- 
polation. But as, up to the present, heating the solution 
beyond 30° C. has not been a success, this temperature will 
be assumed for the purpose of calculation. The effect of the 
current itself is found to maintain the solution at this tern- 



188 



LEAD REFINING BY ELECTROLYSIS. 



perature the year round, the buildings being heated in winter. 
Electrode separation is taken as If inches, which is permissi- 
ble in practice. The figures in Table 72, are not intended 
to give the total power cost, but only that part of it which 
varies with variation in current density. The other losses of 
power, as copper losses and contact losses, should be taken as 
constant for all current densities, for these losses do not depend 
on current density, but on other independent matters, as cost 
of copper for bus bars, and cost of labor cleaning contacts, 
the economical balance for these items being about the same 
regardless of current density. Power is taken at $50 per 
E.H.P. year. 

TABLE 72. 



Current Density 
Amperes per 
Square Foot. 



Volts from Re- 
sistance of 
Solution . 



Polarization 
Volts. 



Total Volts. 



Power Cost at 
95 °c Efficiency. 



10 

12.5 

15 

17.5 

20 



.164 
.205 
.246 
.288 
.328 



.02 
.02 
.02 
.02 
.02 



.184 
.225 
.266 
.308 
.348 



SO . 352 
0.4.30 
0.504 
0.590 
0.665 



The actual total power cost for depositing is about 
SO. 18 higher on account of losses in conductors and con- 
tacts. 

From Fig. 24, taking the 30° C. curve, the use of a solu- 
tion containing 10 gr. lead and 20 gr. SiF 6 per 100 cc. 
would reduce the resistance from 1.35 ohms per inch unit in 
the above case to about 1.05 ohms, when the power cost would 
be somewhat less, particularly for the higher current densi- 
ties, as follows: 



(HOICK OF CONSTANTS. 
TABLE 73. 



189 



Current Density 
Amperes pei 
Square Foot. 



10 
12.5 

1 7 5 
20 



Volts fiom Re- 
sistance of 
Solution. 



'olarization 
Volts. 



0. L28 
0.159 

0.191 
. 224 
. 255 



0.02 
. 02 
. 02 
0.02 

. 02 



Total Volts. 



Power Cost 95% 
Efficiency. 



1 IS 

0.179 

0.211 
. 244 
0.275 



SO. 280 

. 350 
0.392 
0.453 

0.511 



Even stronger solutions than this I have used in 500-lb. 
runs, but the strongest solution yet used in practice contains 
about 17 gr. SiF 6 per 100 cc. 

Our final comparison will take into account power cost, 
depreciation of tanks and interest on metal, the other ele- 
ments entering into the question being small as far as is known, 
and mav be assumed to neutralize each other: 



TABLE 74. 



Current 


Tank De- 
preciation. 


Interest on 
Building, 
Difference. 


Interest on 
Metal. 


Power Cost. 


Total. 


Density. 


A 


B 


A 


B 


10 

12.5 

15 

17.5 
20 


$0,143 

0.114 
0.095 
0.081 
0.071 


SO . 030 
0.023 
0.015 
0.008 
0.000 


$0,423 
0.355 
0.310 
0.278 
0.255 


$0,352 
. 430 
0.504 
. 590 
0.665 


$0 . 289 
0.350 
0.392 
0.453 
0.511 


$0,948 
0.922 
. 924 
0.957 
0.991 


$0,885 
0.842 
0.812 
0.820 
0.837 



While there is not much to choose, the cheapest current 
density is about 15 amperes per square foot. In view, 
though, of the slight differences in operating cost, the choice 
of current density will then be largely influenced by other 
factors, as first cost of plant and elasticity of tonnage treated 
with the plant. 



190 



LEAD REFINING BY ELECTROLYSIS. 



From the standpoint of first cost of plant, on one hand 
we can increase the tank capacity and cut down the size of 
the power plant, and on the other, by increasing the power 
plant we can cut down the cost of the tank plant. I will assume 
that, per ton refined per hour, the power plant must furnish 
23.6 K.W. to overcome contact and other metallic resistance 
anyway ( = .1 volt per tank average), and a variable amount 
of power depending on the current density as follows: 







TABLE 


75. 




10 


45.2 


K.W. 




68.8 total K.W 


12.5 


55.2 


1 1 




78.8 " 


15 


65.4 


1 1 




90.0 " 


17.5 


75.8 


i t 




99.4 " 


20 


85.8 


1 1 




109.4 " 



The cost of power plant may be roughly taken as $135 
per K.W., and of the tank plant, disregarding handling ma- 
chinery and copper bus bars as practically constants, but includ- 
ing tanks, electrolyte, and floor area, at $15,000 per ton per 
hour, with a current density of 12.5 amperes. 

Cost of variable items in plant for solution with 17 gr. 
SiF 6 per 100 cc: 

TABLE 76. 



Current Density. Power Plant. 


Tank Plant. 


Total. 


10 

12.5 
15 
17.5 

20 


$ 9,270 
10,650 
12,150 
13,400 
14,750 


$18,750 

15,000 

12,000 

10,714 

9,37', 


$28,020 
25,650 
24,150 
24,114 
24,125 



With the stronger solution, 20 gr. SiF 6 and 10 gr. Pb 
per 100 cc. 





CHOICE OF CONSTANTS. 
TABLE 77. 


191 


Current Density. 


Power Plant. 


Tank Plant. 


Total. 


10 

12.5 

15 

17.5 

20 


$7,960 

9,050 

9,370 

10,350 

11,250 


$20,000 
16,000 
13,333 
11,428 
10,000 


$27,960 
25,050 
22,703 
21,778 
21,250 



From these results it appears that in future progress will 
tend to higher current densities and stronger solutions, and 
will reach probably 20 amperes per square foot, if no unfore- 
seen objection crops up. 

A plant built for 15 amperes per square foot can be arranged 
as to be able to stand a 33% overload if the solution is 
strengthened up somewhat. 

No combination of current density and solution strength 
should be used, at which a solid lead deposit may not be 
obtained, otherwise the increased acid loss would offset any 
advantage gained. 

Choice of slime process. — Out of a large number of slime 
processes described in more or less detail in Chapter II, the 
following only will be considered as being available at the 
present time for practical work, the others being too little 
developed or to apply only to special cases. The sodium 
sulphide process has been given an extensive trial at Trail, 
but full particulars have not yet been given out. The pro- 
cesses considered below have been the subject of much experi- 
ment and have been or will probably be used in practical 
work. 

(la) Melting with sulphur to matte and slag, especially 
for slime containing little or no bismuth. 



192 LEAD REFINING BY ELECTROLYSIS. 

(16) Melting to clore, matte, and slag, especially for slime 
containing bismuth. 

(2) Extraction of copper and arsenic in sulphuric acid 
solution and antimony from the residue with hydrofluoric acid. 
Oxidation to be secured by (c) roasting with sulphuric acid, 
(d) drying in air, and (c) by ferric sulphate produced elec- 
trolytically. 

Process (la) converts the copper and silver present into 
cuprous and silver sulphides, which have to be reduced to 
metal and electrolytically refined. All methods of converting 
the matte into metal so far successful are fairly expensive, 
and for that reason this process shows up at a disadvantage 
compared with the others, when the amount of silver and 
copper are at all large. But for slime containing only a little 
copper and silver this process will do excellently. Assuming 
that the lead bullion being refined contains 50 ozs. silver per 
ton and 0.2% copper, beside 20 lbs. of antimony and 8 lbs. 
of arsenic, the costs would be about as follows, on a large scale 
of say 100 tons of lead per day: 

TABLE 78. 

Per Ton 

Lead. 

Drying and oxidizing slime (Fig. 57) SO . 05 

Melting in iron pot, coal, labor, repairs (Fig. 58) 0.08 

Sulphur, 2 lbs . 03 

Grinding matte, heating with HsSO^ and melting. Coal, labor, re- 
pairs, HsSO* 0.11 

Electrolytically refining alloy for copper 0.04 

Melting and refining silver 0.11 

Grinding and leaching slag for SbF 3 solution, and smelting and refining 

PbS0 4 residue 0. 12 

Depositing 20 lbs. antimony . 40 

SO. 94 



CHOICE OF CONSTANTS. 193 

Process (16). For lead bullion containing 20 lbs. antimony, 
2 lbs. copper, 6 lbs. arsenic, 2 lbs. bismuth, and 70 ozs. silver 
per ton, the costs are calculated to be about as follows: 

TABLE 79. 

Per Ton 

Lead. 

Drying and oxidizing slime $0 . 05 

Melting in iron pot 0.08 

Treating matte as above . 05 

Electrolytically refining dore, and recovering bismuth, copper, silver, 

and gold 0.21 

Grinding and leaching slag as above 0.12 

Depositing 20 lbs. of antimony . 40 

$0.91 

Usually slime will contain too much copper for the above 
processes, and this would be more certain in future for elec- 
trolytic refineries, as I will attempt to show. Very often 
lead bullion as it flows from the lead-furnace contains copper. 
As the smelter has not been paid anything for copper in his 
bullion, and can get something, though nowhere near its 
full value, if in matte, usually the lead bullion is cooled and 
skimmed in the lead " cooler," the dross with or without 
liquating as much lead as possible, going back to the blast- 
furnace, and the copper being eventually recovered as a matte 
containing probably on the average 40% of copper and 10-15% 
lead. The lead in this matte is not usually paid for or re- 
covered, and counting in the lead loss, the copper of the dross 
has been reduced in value by as much as 6 cents per pound. 
In refining bullion by the Parkes process, the refiner has no 
advantage over the smelter in recovering copper, as the refinery 
also puts the dross through a lead-furnace. An electrolytic 
refinery is, however, free from these objections, and certainly 



194 LEAD REFINING BY ELECTROLYSIS. 

when the smelter and refinery are under the same control the 
practice of skimming off as much dross as possible will be found 
uneconomical. Custom refineries using the electrolytic pro- 
cesses are in a position to credit the smelter enough for cop- 
per in bullion to discourage the skimming process. For these 
reasons it may be expected that the tendency will be toward 
more copper in the bullion, and not less, so that the consid- 
eration of bullion carrying more copper, from h to 1% or more, 
is important. 

Probable cost of treating slime by (2c), from lead bullion 
containing per ton 20 lbs. antimony, 10 lbs. copper, 5 lbs. 
arsenic, and 70 ozs. of silver, beside which 5 lbs. of lead remain 
in the slime: 

TABLE 80. 

Per Ton 

Lead. 

Sulphuric acid lost, 15 lbs SO . 12 

Hydrofluoric acid lost . 08 

Operation electrolytic tanks, including power at $50 . 48 

Power, 140 K.W $25.60 

Labor 7.50 

Repairs 2.50 

New anodes 12 . 20 

Per day (100 tons) $47.80 

Repairs and supplies . 10 

Melting and refining dore at \ cent per oz 0.14 

Labor not already included 0.18 

Coal 0.05 

SI. 15 

If bismuth is present it will be recovered from the sedi- 
ments deposited from the sulphate solution, and from the dore 
bullion, at small cost. 



CHOICE OF CONSTANTS. 195 

(2d) This process gives a similar result and is a little in- 
ferior, though the loss of sulphuric acid is less. On the other 
hand the roasted product is not so readily leached, and some 
sodium nitrate is required to finish the oxidation. 

(2c) Ferric sulphate process, for same bullion as assumed 
for (2c): 

TABLE 81. 

Per Ton 

Load. 

Operation of elect rolytic tanks, including power at $50 per year. Cop- 
per tanks operate at 1.75 volts and antimony tanks at 2.9 volts, 

with 90% efficiency each SO. 63 

Power, 175 K.W $32 . 40 

Labor 10.00 

Repairs 12 . 50 

New anodes 8 . 00 

Per day (100 tons) $62.90 



Hydrofluoric acid loss, 1 lb. at 7.5 cents 0.08 

Sulphuric acid, 10 lbs 0.08 

Melting and refining dore at A cent 0.14 

Labor not already included . 24 

Repairs and supplies 0.10 

Coal for melting, roasting matte, melting antimony . 03 

$1.30 



Credit for 30 lbs. copper recovered from matte 1 . 05 

Cost $0.25 

If bismuth is present, it will be recovered in a special bullion 
from smelting the leached matte, and to a small extent in 
the dore bullion. 

The above results do not include superintendence and assay- 
ing, metal losses, or interest on plant, but the comparison is 
still valid. 



196 LEAD REFINING BY ELECTROLYSIS. 

In a general way, the ferric sulphate method, beside being 
the most economical in net operating cost, is the cleanest, 
easiest, and quickest, and will cause less loss of precious metals 
through the various channels of loss, for the slime is not dried 
at all until the final melting to dore bullion. The ferric sul- 
phate method will recover all the precious metal values shown 
by corrected fire assay, or more. 



CHAPTER VII. 

REFINERY CONSTRUCTION, OPERATION, AND REFINING 

COSTS. 

The general arrangement of most, if not all, large elec- 
trolytic copper refineries is on the one level plan with indus- 
trial railway running between the different departments, motive 
power being generally provided by electric locomotives. This 
arrangement can be safely copied for electrolytic lead refineries. 
In the casting plant the molds to receive melted metal may 
be on a level a few feet below the general level; but in stacking 
the anodes, ready to be carried off by the tank-load by the 
electric cranes, they can be easily hoisted the necessary few 
feet. 

In considering the level question, the tank-room can be 
regarded as receiving and delivering material on the same 
level. 

The melting plant should be so situated and arranged 
that lead and bullion may be handled from and to the railroad 
cars as simply as possible. In order to deliver the cast anodes 
and pig lead on nearly the same level as the tank-house and 
shipping track, it is preferable to have the melting- furnaces 
at a higher level, so that the metal may flow by gravity through 
siphons to the molds. If a Rosing steam pump is used, the 
pots may, of course, be brought down to the same level; but 

197 



198 



LEAD REFINING BY ELECTROLYSIS. 



this has been tried at Trail and I believe not found entirely 
satisfactory* 

The melting of the bullion bars and anode scrap and of 
the cathodes has been done up to the present by simple melt- 
ing down in kettles. The cathodes are wet when they come 
from the refinery and have to be dried before coming into con- 
tact with melted lead in the pot. The usual plan is to pile 
the cathodes high above the pot and melt down slowly. A 
cover and pipe should be provided over the kettle to carry 
off fumes. When the lead is melted finally about 4% of dross 
floats on top, which is skimmed off by hand, a slow and labori- 
ous method. A Howard skimmer, such as used in the 
Parkes process, to take off the dross, would be a desirable 
adjunct. Plate 4 shows the Trail melting plant, 1903-1904. 

The dross ordinarily produced is less pure than the lead 
and contains more silver. Table 82 shows the comparative 
analyses of lead and dross from the same meltings at Trail. 

TABLE 82. 





Fe 


Cu 


As 


Sb 


Zn 


Ag. Ozs 


Lead 


.0010% 
.0016% 

.0008% 
• 0011% 


.0003% 
.0005% 

.0009% 
.0010% 


.0002% 
.0003% 

.0001% 
.0008% 


. 0010% 
.0016% 

. 0009% 
.0107% 


None 




Dross 




Lead 


.24 


Dross 









I believe that the present method of melting the cathodes 
is capable of considerable improvement, along the line of sav- 
ing labor, and making little or no dross. The piling of the 
cathodes above the pot, and the necessity of steering them 



* See, however, description of lead pumps in Appendix I. 




PLATE 4. 
Truswell Anode Mold and Anode. 



page 199 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 201 

into the pot properly as the charge settles down, requires some 
labor, as does the handling of the dross. What is wanted is 
to dump the damp cathodes by the carload and have no more 
labor involved until the lead is cast. This end could be 
achieved by dumping the cathodes through the roof of a 
preheating reverberatory furnace at a level just above that 
of the refined-lead kettle. The reverberatory could have a 
cast-iron or steel bottom to prevent its being broken up by 
the falling lead. 

A very small heat supply will suffice to melt lead in this 
way, and if the furnace gases were kept reducing at the same 
time, little or no dross would be formed from oxidation of the 
melting cathodes. At 50% heating efficiency, which does not 
seem high for a furnace working at the melting-point of lead, 
8 lbs. of coal would be theoretically required to melt a ton 
of lead. The waste heat from the bullion or refined-lead ket- 
tles could be applied veiy easily to the melting in the rever- 
beratory. A similar operation is the liquation of bullion in 
a reverberatory, to soften it for the Parkes process, which 
requires with a 35-ton furnace 24 lbs. of coal per ton of lead 
melted.* The objection might be raised that the resulting 
lead will be slightly less pure, which is undoubtedly a fact. 
Electrolytically refined lead usually contains from .1 to .5 ozs. 
of silver, which is in practical work almost entirely due to 
slime not washed off the surface of the cathodes. In taking 
a crop of cathodes from a tank, the disturbance of the tank 
or unavoidable contact of the anodes and cathodes, is apt 
to loosen some slime from the anodes and get part of it on 
the cathodes. When these are dipped in muddy wash-water, as 

* Collins, "The Metallurgy of Lead," page 288. 



202 LEAD REFINING BY ELECTROLYSIS. 

is sometimes clone, the result is an even distribution of part of 
the slime over the surface. In the ordinary melting quite a 
little of this slime goes into the dross, as will be seen from 
Table 83, from the United States Metals Refining Company. 

TABLE 83. 

Lead 25 ozs. per ton. 

Dross 20 mesh oversize 1 . 836 " " " 

" 40 ■ • 1.776 " " " 

11 60 " " 1.75 " " " 

" though 60 mesh 3.66 " " " 

Assuming; \°7 r of dross reduced and melted into the lead, 



L .r- 



the lead would have carried approximately .35 ozs. instead 
of .25 ozs. of silver, provided all the adherent slime was taken 
up by the lead, which is not probable, as some of the slime would 
probably remain in the furnace as dross. The difference in 
the amount of the other impurities in the lead would be too 
small to be noticeable. 

For a combined smelting and refining works, the lead 
should be cast from the blast-furnaces into anodes direct. 

In refineries, anodes are cast in open molds lying in a semi- 
circle in front of the pot, to which the usual lead launder 
reaches from the discharge end of the siphon or pump. These 
may be seen mounted on a long car in the photograph of the 
Trail melting plant of several years ago, Plate 5. 

The use of a rotating table, on which the molds are placed, 
similarly to the casting machinery used in electrolytic copper 
works, has been proposed, but it is doubtful if it would save 
anything in cost of casting. It is quite possible that the 
adoption of rotating molds in casting copper anodes was a 
necessity, because copper could not be conveniently run through 
a long launder to a semicircle of molds, and copper requires 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 203 

so much cooling that it is necessary to pass the casts under 
water, so the successful use of rotating molds in copper cast- 
ing is no valid argument for its adoption in lead casting. 

An anode mold is shown in Fig. 29. The following re- 
marks will be of value in designing them. No draft is pro- 
vided at the under side of the lug. The sides of the lug 
should not be too steep, as the anode in contracting draws 




•ly*- 



v 



%-, 



Fig. 29. 



the lugs against the mold at those points, making the anode 
stick in the mold so that it has to be forced out. Except on 
the under side of the lugs, which are required flat to make the 
anode hang straight in the tank, a good draft may as well be 
provided as not, to facilitate the removal of the anodes with- 
out its being necessary to pry them out. 

No trouble from sticking is experienced at the under side 
of the lugs, as the anode contracts somewhat after it has solidi- 



204 LEAD REFINING BY ELECTROLYSIS. 

fied. Once in a while the anode molds should be sprinkled 
with clayey water, which rapidly dries off the hot iron, and 
leaves a coating to which the lead cannot stick. The block 
at the top between the two lugs is separate and removable, 
and gives a place to put a bar in to lift the anode slightly from 
the mold, so that it may be engaged by hooks. A com- 
pressed-air hoist on a light jib-crane enables one man to lift 
anodes and stack them rapidly. 

The anodes molded in the way mentioned suffer from 
irregularities of form or weight, as would naturally be ex- 
pected when the workman's sole means of judging the amount 
of metal run into each mold is by eye. Then the molds and 
inflowing bullion are at various accidental temperatures, so 
there can be no uniform procedure for getting the right amount 
of metal in the lugs and all the different portions of the plates. 
Even if the mold is perfectly level an anode may average 
considerably thicker on one end than the other from the unequal 
flow and chilling of the lead. When these irregularly-formed 
plates are suspended in the tanks the thinnest plates will be 
entirely dissolved, while considerable metal remains on some 
of the others, increasing the proportion of anode scrap to be 
remelted. A stiffening rib about \ inch deep and 2 inches 
wide is usually cast across the top of the anodes to prevent 
their collapsing and falling in the tank at the end of the run 
when the metal has been nearly all dissolved. Part of the 
rib, of course, remains when all of the lower part of the 
anode has been decomposed. 

For the best work and highest efficiency in the tanks, the 
anodes should all be of the same weight, and slightly thicker 
at the top than at the bottom. For these reasons the idea 
of using closed molds for casting the anodes has been an attrac- 




« 

H 

CO ^ 

§ 4 



^ .s 




PLATE 6 

Truswell Anode Mold 



page 207 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 209 

tive one. The closed molds should aim preferably to cast the 
anodes bottom up so that the dross rising in the cooling liquid 
metal can not flow into the lugs, both weakening them and 
sending the impurities back to the melting-kettle in greater 
relative amount than they exist in the bullion. With anodes 
cast in closed molds experimentally, the tank efficiency has 
been raised at Trail to 95%,* as against the usual 90%. Mr. R. 
Truswell, Trail, B. C, has applied for United States,! Canadian, 
and English patents for his anode-mold, which is shown in 
the photographs supplied by Mr. Truswell, Plates 5 and 6. 
The following description is from Mr. TruswelFs article :{ 

"The illustrations show a new mold that I have devel- 
oped for the purpose of casting anodes; it has the advan- 
tage that the plate, being enclosed during the process of cast- 
ing, will be of uniform thickness and not liable to be warped 
or twisted. To prevent the dross or spongy characteristics 
found in some plates it is cast on end, and under a head of 
fluid metal to ensure its soundness. The dross rises toward 
the gate, and, as this is near the lower end of the plate, its 
defects are less noticeable than when the method of pouring 
is not that specified. 

"In the illustrations, Figs. 30 and 31, are profile views of 
the plates for which the construction of the mold is adapted. 
Fig. 32 is a front elevation of the mold mounted as for pour- 
ing. Fig. 33 shows a cross-section on AA in Fig. 32, and 
Fig. 34 is an end elevation of the mold, inverted after cast- 
ing and with the mold opened for removal of the cast plate. 
Fig. 35 is a detailed cross-section of the slide and nut of 



* Communicated, Mr. W. H. Aldridge. 

f United States Patent, 823977. June 19, 1906. 

J Engineering and Mining Journal. May 5, 1906. 



210 LEAD REFINING BY ELECTROLYSIS. 

the opening portion of the mold, and Fig. 36 is a detail 
section showing a modified form of joint between the head 
and plate portion of the mold, by which the parts which en- 
close the head are sustained when inverted. 

"The anode plate is represented by 2; that shown in Fig. 
1 is provided with laterally projecting horns, 3, by which the 
plate is supported on the walls of the tank; that shown in 
Fig. 2 has eyes, 4, which are usually bent and cast into the 
plate, but may be cast with the plate. 

"The main body or plate portion of the mold is formed 
of two recessed parts, 5, secured by any suitable fastening, 
so that the recesses when together will leave a space 6, to 
form the mold of the lower or uniform portion of the plate. 
These recesses are carried to the end of the mold, so that the 
metal may be poured from that end which forms the lower 
part of the plate when in position in the tank. 

"The head of the plate, including the horns, 3, or eyes, 4, 
as the case may be, is formed by two portions, 7, each hav- 
ing recesses, 8, to form the desired width and shape of head 
which they entirely enclose. These portions, 7, are slidable 
outward from the middle plane of the mold a sufficient dis- 
tance to clear the mold from the projecting portions of the 
plate which has been cast within it. The contiguous edges 
are beveled as at, 8, so that they will form a close joint 
together. 

" Each drawback portion, 7, is furnished on each side with 
outwardly projecting members, 9, by which they are sup- 
ported on V-shaped slides, 10, on brackets, 11. These project 
from the adjacent sides of 5, and the, parts. 7. are slidable to 
or from the plane of division by screws, 12, having right- and 
left-hand threads on their opposite ends. These pass respect- 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 211 



Fig. 30. 



Fig. 31. 



!•! 



«>16 



SI,— •' 8 




Fig. 32. 



Fig. 33. 




Fig. 34. 



Fig. 35. Fig. 36. 



212 LEAD REFINING BY ELECTROLYSIS. 

ively through correponcling nuts, 13, secured by screws to 
the parts, 7. The screws, 12, are supported in bearings, 14, 
upwardly projecting from the outer ends of the brackets, 11, 
and are rotatable therein by a hand-wheel or crank, 17, on 
either one. A shaft, 15, extended between the screws and 
connected to them by beveled pinions, 16, enables them to be 
simultaneously operated. 

"The mold is pivot ally mounted by trunnions, 20, secured 
to or forming a part of the plate portion, 5, in a frame, 21, 
provided with wheels; it is furnished with a hand- wheel, 22, 
by which it may be inverted in the frame. 

' ' In operation the parts, 7, are tightly closed and the mold 
is inverted to bring the open end of it uppermost as in Figs. 
32 and 33. The metal is then poured in, and when set the 
mold is again inverted, the parts, 7, withdrawn, as represented 
in Fig. 34, and the plate drawn from the mold. Some draft 
may be desirable in the width and thickness of the plate, to- 
wards the open or pouring end of the mold, to facilitate its 
removal. It may also be necessary to support directly the 
head-end members, 7, when the head mold is closed, to enable 
them to sustain their weight and that of the fluid metal within 
the mold. For this purpose some such modification as is 
shown in Fig. 36 may be adopted. In this engaging lips, 18, 
are provided on the contiguous edges of 5 and 7, the lips secu- 
ring these parts of the mold against separation endwise. 

"These molds can be made with water-jackets, and can 
be mounted on cars in any number desired, and can all be 
opened at once by the turning of one lever." 

The molding of the refined lead calls for no special remarks, 
the usual method described by the authorities on lead smelt- 
ing being used. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 213 

The sampling of bullion bars received to be refined, would 
be done in the ordinary manner, taking five punches diagonally 
across a row of five bars on the top, and then turning them 
over on the bottom and taking one punch each from each 
bar in the same manner, but on the opposite diagonal.* Sam- 
pling anodes does not give the same result as is got from the 
bars from which the anodes were cast, and both are lower 
than a dip sample taken from the lead flowing to the molds. 
I believe the sample taken by punching anodes in a small 
number of places can be relied on very closely, for usually 
in casting the lead chills in the anode mold pretty quickly, 
especially in the lugs and corners, and you have a plate of com- 
paratively small and even thickness to punch several times 
on both sides. The following results in Table 84 are from a 
test made at Trail, July, 1902. 

TABLE 84. 



6 

01 


Bar 

Sample 


Dip Sample 
Anode. 


Punch from 
Top Anode. 


Punch from 
Bottom Anode. 


Average Top 
and Bottom. 


S 

d 


Au 


Ag 


Au 


Ag 


Au 


Ag 


Au 


Ag 


Au 


Ag 


1 j 2.79 

2 2.91 


322.2 
331 . 1 


2.88 
2.92 


328.6 
333.6 


2.72 
2.90 


316.0 
329.6 


2.84 
2.99 


324.7 
331.6 


2.78 
2.94 


320.7 
330.6 



Size of tanks. — This will have to depend on the capacity 
of the plant. For fair-sized or large plants a current of 3500 
to 6000 amperes is satisfactory, the latter probably being 
more favorable. It seems evident that the larger the tanks 
can be made the smaller the cost of tank construction and 



*Hofman, "Metallurgy of Lead," 1899, page 351. 



214 LEAD- REFINING BY ELECTROLYSIS. 

maintenance per ton produced. The separation of electrodes, 
which is usually expressed in distance from center to center 
of anodes, varies from 4f to 4 1 1 - o y as follows: 

TABLE 85. 

Trail, B. C 4.375 inches. 

Grasselli, Ind 4.625 " 

Newcastle-on-Tyne 4.15 " 

The anodes are about 3 to 4 inches narrower than the 
tanks themselves. The anode width is usually about 2 feet, 
but this can be increased as well as not to 2 feet 6 inches, 
or even 3 feet, and thereby shorten the tank and reduce 
the number of electrodes to be handled. 

TABLE 86. 

Plant. Anode Width. Tank Width. 

Trail, B. C 26 inches 30 inches 

Grasselli, Ind 24 " 30 " probably. 

Newcastle-on-Tyne ... 33 " 37 ' ' 

The depth of the anode exposed to the electrolyte is from 
2 feet 10 inches to 3 feet. 

TABLE 87. 

Plant. Anode Depth. 

Trail, B. C 34 . 5 inches. 

Grasselli, Ind 36 

Newcastle-on-Tyne 34 

Adopting the maximum dimensions now used in each case 
for a 6000-ampere current, anodes 33 inches wide and 36 inches 
deep, current density 17.5 amperes per square foot, space 
center to center of anodes 4 h inches, the tank would need 
the following dimensions inside: length 8 feet, depth 3 feet 
9 inches, width 3 feet 1 inch. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 215 

Wood tanks have been used exclusively, for which South- 
ern yellow pine is good material. Cedar was tried at Trail 
first and found rather soft, and more recently better results 
have been obtained with fir. 

The tank of the future, in my opinion, will be made of 
reinforced concrete, saturated with sulphur by immersion 
in a sulphur bath.* They are cheaper than wood and ab- 
solutely acid proof. A small tank of this kind in my labor- 
atory containing hydrochloric acid is in the same condition 
after several months' standing. There is absolutely no 
action of the acid on the tank. Other tanks are now 
being tested on a practical scale, with good results, as far 
as corrosion goes, though they cracked somewhat in the 
corners. 

The preparation of a concrete tank of any particular size 
and shape, reinforced or not, does not need any very long 
description here. An article by Mr. D. H. Browne f gives 
a good description of the manufacture of electrolytic tanks, 
from which the following remarks are taken: 

"The first requisite is a good, slow-setting cement. Slow- 
ness of set is necessary because in building large tanks it re- 
quires ten or twelve hours or even more to carry up the walls 
to the required height, and as the ramming must be contin- 
uous throughout this entire time, it is evident that if the bot- 
tom took its initial set before the sides were completed it would 
be injured by the vibration. The cement, therefore, should 
take longer to set than the tank to complete. Cheap cement 
is worse than useless. Savior's cement has proved a reliable 



* Patent applied for. 

f " Electrochemical and Metallurgical Industry," Vol. I, page 273. 



216 LEAD REFINING BY ELECTROLYSIS. 

article, but any brand which will stand the 'pat' test will 
be satisfactory. 

"The 'pat' test is made by mixing a handful with water 
to a stiff paste and working the same on a glass plate into 
a cake about half an inch high and 3 or 4 inches in diameter. 
The surface should be troweled smooth and the sides brought 
down to a thin edge. This is allowed to stand a few hours, 
then is covered with a wet cloth and set aside in a cool place 
over night. If it sets slowly and shows no cracks on the sur- 
face or at the edges it will answer. 

"For the best work crushed granite should be used. 
This has a rough granular fracture or 'bite/ into which the 
sand and cement lock better than with any other rock. As 
the stone used is the weakest part, and as a good concrete, 
when broken, shows fracture across, and not around the par- 
ticles of stone, it is important to use the best rock available. 
Failing granite a trap rock or blue diorite is a good substitute. 
The size of the rock depends on the thickness of the walls: 
a safe rule being that no piece should be over one-quarter 
the thickness of the wall in which it is used. For ordinary 
tanks material passing through a screen of 1J inches and over 
a screen of one half inch is satisfactory. The material smaller 
than one half inch should be rejected, as it interferes with the 
filling of the voids. 

"The solidity of concrete depends largely on the care with 
which these voids are filled. To determine the void space, 
take a pail of crushed rock, calculate the volume and find 
the weight. Add now water till the pail is full and weigh 
again. Calculate the volume of the water and simple pro- 
portion shows the empty space between the particles of rock. 
This space must be filled with sand, of which in turn the voids 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 217 

must be filled with cement.* The voids of cement are in 
their turn filled by the water absorbed. Hence for strong 
concrete the common use of the formula, '4 parts rock, 2 parts 
Band, 1 part cement. 1 For less careful work a larger pro- 
portion of rock is often used. 

"To mix the cement a tight mortar-box or floor and a 
measure holding one cubic foot are needed. The rock should 
be thoroughly washed and the sand screened from clay or 
gravel. One cubic foot of cement to two of sand is mixed 
on the dry floor to an even composition, and to this four cu- 
bic feet of stone are added, and the mass thoroughly shoveled 
over. Water is now added, so that, while no muddiness is 
apparent, each particle is moist. The mass is again shoveled 
over and is now ready for the mold. , 

"This mold may be of any shape whatever. It is set on 
a solid floor, with a sheet of building paper underneath so 
that the tank does not bind to the floor. The sketch shows 
a form for a commercial-plating bath. The outer frame is 
trued at right angles and braced by struts to the floor to pre- 
vent bulging of the sides under the rammer. The concrete 
is now shoveled in, a few inches at a time, and thoroughly 
rammed until water shows at the surface. For a tank of the 
size shown three men are needed ramming and two men mix- 
ing and handling concrete. The tools needed are iron rammers, 
about 2 inches thick and 3 or 4 inches square, with a sleeve 
for a wooden handle. Such a tool, handled with a short, 
stiff blow, is better than a lighter tool, with a springy blow, 
the idea being simply to drive out the air from between the 
particles and completely fill the voids. 

* This method of test not now considered entirely correct. 



218 LEAD REFINING BY ELECTROLYSIS. 

"As soon as the bottom is of the desired thickness the 
inner frame is put in place and braced by cross-pieces to pre- 
vent inward bulging. The sides are now rammed up, a few 
inches at a time. It is not desirable to lay the sides in lay- 
ers, but rather to carry them up without coursing or strati- 
fication. One thing is very important — that there be no stop- 
pages. If a mealtime intervenes the men should be relieved 
one at a time, so that no pause occurs till the tank is 
completed. 

' l The top finish is put on by bringing the concrete to within 
a quarter of an inch from the top of the mold and carry- 
ing this up with equal parts sand and cement troweled to a 
smooth surface. Any openings or holes in the tank wall 
are made by inserting a block of wood of the desired size in 
the side walls. After the tank is set the wood can be drilled 
or broken out. 

"Three or four days should elapse before the moldboards 
are taken down. The inner frame is removed by unscrew- 
ing the angle irons shown, when the side boards will drop in- 
ward without any difficulty. The outer form falls apart on 
removal of the the rods. If necessary the inner surface can 
be finished with a coat of sand and cement, but if planed 
boards were used for the molds the surface is usually quite 
smooth. 

" Concrete will not stand strong acids; caustic or chlorine 
has no effect upon it. A coating of paraffine or tar would 
help it to resist acids. It should not be subjected to sudden 
changes of temperature. If the heat be brought up gradu- 
ally it will stand fire. It can be handled or lifted like a block 
of granite if ordinary care be used to prevent the tools from 
bearing against the sharp edges of the tank. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS- 219 






A tank with 6-inch bottom and 4-inch sides, containing 
24 cubic feel of concrete, can be set up and completed by 
live men in one day. The cost decreases with the number 
of tanks built at one time and the facilities for handling con- 
crete. Building four tanks of this size per day the cost per 
tank was as follows: 

TABLE 88. 

Carpenter and blacksmith labor on molds $1 . 75 

Concrete work labor, 30 hours at 17£ cents 5.20 

3.5 cubic feet cement at 60 cents 2. 10 

7 cubic feet sand .25 

14 cubic feet crushed trap rock 3 . 00 

$12.30 

"Including finishing, taking down molds, cementing in 
rubber-pipe connections, about SI 5 will cover the cost of 
building a tank as above described, the dimensions of 
which are about 3 feet wide, 9 feet long and 2 feet deep. 
No construction of lead, slate or wood can be made 
which will fulfill all the requirements of the case for this 
sum." 

To make the tank acid-proof, after standing moist for 
several weeks until well set, the tank is dried out pretty well, 
and then lowered into an iron vessel containing just-melted 
sulphur. The sulphur is gradually heated to 150° C. or so, 
but not to the thickening point. This should take quite a 
number of hours, perhaps 12, steam coming off regularly as 
long as the temperature is rising, and of course removing 
with it all permanent gases present in the concrete. The 
sulphur is then allowed to cool slowly during another 6 to 12 
hours, when the sulphur penetrates the crevices and cracks 
in the concrete. Probably the atmospheric pressure helps 



220 LEAD REFINING BY ELECTROLYSIS. 

in this, as the reabsorption and contraction of steam in cooling 
would make a vacuum in the concrete. 

The tank is then lifted out, and after cooling to perhaps 
80° to 90° C, is quickly dipped in again and taken out. This 
chills a thin, smooth layer of sulphur on the tank, and fills 
any cracks, while the coating produced does not come off or 
crack off during long periods. A coating of asphaltum paint 
applied later is partially absorbed by the spaces between the 
sulphur crystals and adheres very well. In fact the surface 
of the finished tank will soak up quite a little paint, melted 
paraffine, etc. 

Wooden tanks have been built with all bolts clear of the 
wood and with bolts through the wood. Figs. 37 and 38 show 
the two methods of construction. The bolts in the wood 
show corrosion badly sometimes, especially where the bolts 
pass from one plank to another. Iron is not rapidly attacked 
by the lead-depositing electrolyte when no current is passing 
in the neighborhood, but as there are slight differences of 
e.m.f. between different parts of the tank, it would only be 
expected that lead would deposit on one part of an iron 
conductor touching the solution at several places, and iron 
dissolve at another. For this reason bolts clear of the wood 
could be expected to last longest. The first tanks at Trail were 
made in this way, but in the tanks used now the bolts pass 
directly through the wood. If the bolts could be surrounded 
by a rubber tube, or copper bolts used, they would then be 
most successful. The expedient of pouring hot paraffine, 
pitch, etc., through the holes before putting in the bolts, seems 
to help a little, but not to be entirely successful. In design- 
ing a wooden tank, the placing of the bolts should be studied 
not only from the mechanical standpoint, but to reduce elec- 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS- 221 

trolytic corrosion of the iron as far as possible, which I am 
satisfied is the main cause of the failure of the bolts. 

As an example, if two tanks are bolted together, as shown 
in Fig. 39, it is evident that the current will tend to pass 



i 


i E2 


! 

°i } 

i i 

• 


I'j i E3 


■ 

i 


!ji ! s 


Q 


III- 


© (JU) (£||) (^ — 


^ D . 




Fig. 37. 



through the wet wood to the iron bolt in one tank, depositing 
lead on it probably, and pass from the bolt to the solution 
in the other tank, with too rapid corrosion of the bolt. It 
would be expected that the greatest effect on the bolt would 



222 



LEAD REFINING BY ELECTROLYSIS. 




_ 



.1. 



'H 



/i 


-! 5 


*u_ 




i 


> 

-- 


•J)> 5? '^3 j^j «# if-'i ,f| 


ft 

-- ] 

1 


1> !! }j <f 


{ 


4-1 X 4J.i Ii 4- 


i 
i 


U> JJ jjj j < 


f 


: I> !! < 


3?f 




t 


^H^^-frf-^:=*--1-^=t-= 




^ e # ^ 



Fig. 38. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 223 

be where it passes from one plank to another, unless the joint 
in the wood were perfect. As a mat ten- of fact, that is where 
bolts fail usually. Bolts as shown in Figs. 40 and 41 would 
be apt to carry current as shown, if the tanks rested on a wet 
beam. 

A few tenths of an ampere would cut a bolt through in a 
moderate time. 

Three- or four-inch planking should be used for tank walls 
and bottom, four inches being best. Tanks with four-inch 
sides do not need a bolt across the top in the center or braces 
to hold the sides from bulging. The use of feather and groove 
in the joints is preferred by some and not by others. 

The problems to be settled in connection with the tanks are 
their arrangement and differences of elevation for circula- 
tion purposes. Two general systems of locating tanks for 
the multiple system are in use both in electrolytic copper and 
lead refining. The older method, which we may call the 
11 cascade/' originated, I believe, by Mr. F. A. Thum, uses 
double rows of tanks end to end, each pair at an elevation 
of 2J-3 inches above the next pair, while a continuous cir- 
culation of solution flows from the highest tank at one end 
to the lowest at the other. The newer arrangement pat- 
ented by Mr. A. L. Walker * offers some important advantages 
especially for copper refineries, in requiring less space, less 
copper conductors by far, and saving some power. For lead 
refining, considering that the number of tanks, amount of 
conductors and power are only about a third as great per ton 
produced as in copper refining, it is evident that these advan- 
tages are much reduced when the Walker arrangement is ap- 

* U. S. Patent 687800. December 3, 1901. 



224 



LEAD REFINING BY ELECTROLYSIS. 



EC 






Ik 


















t t.l 

1 1 1 




i I 1 
t t | 






c 


:l- 


------- 


_-=< 


K.- 


_-_-_-- z : 




> 



Fig. 39. 



4 



i \ . u 



3 E 



Ml! 



fe. 



Fig. 40. 




Fig. 41. 




a 



H o 

— i a u 
<1 « pq 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 227 

plied to lead, while the disadvantages, which are of a mechan- 
ical nature due to greater crowding, are increased somewhat. 

There is also more chance for injury to the workmen with the 
Walker system. 

The first tanks at Trail, shown in Plate 4 7, were arranged 
by the cascade 1 system. The next tanks had the newer ar- 
rangement, and the two systems were operated side by side, 
the old arrangement giving much higher efficiency and better 
satisfaction. As thereafter more tanks were added according 
to the old system, it is to be inferred that the old system 
was considered best. The Grasselli plant of the United States 
Metals Refining Company uses the Walker system. 

The Walker system uses less power and less copper bus 
bars, which latter may be taken to be about $50 less per ton 
per day installed in first cost of copper. The saving in space 
would not amount to over about 80 sq. ft. of area per ton of 
lead per day, as the tanks are usually installed; worth say 
S80 per ton refined per day. The power lost in the bus bars 
from tank to tank with the old system, using 5 sq. in. of 
copper for 4000 amperes, is about 85 watts per tank, or per 
ton per day, about 235 watt days = 5.6 K.W. hours per ton 
lead, or 4.3 cents' worth, with power at $50 per year. A loss 
of current efficiency of 6% (which may be expected when 
leaking wooden tanks are placed in two continuous rows close 
together) would offset this gain. With absolutely tight and 
non-conducting concrete tanks mentioned on page 23, there 
would be however no objection from current leaks. The only 
saving by the Walker system is then about $130 per ton per 
day in first cost, and 4 cents per ton in power, a total of about 
8 cents per ton, figuring interest on cost for extra copper and 
extra space as high as 10%. 



228 LEAD REFINING BY ELECTROLYSIS. 

Subdivision of tanks into blocks. — With the cascade sys- 
tem we can have a sloping floor so that the tanks are every- 
where at the same height above the floor, which is however 
not as good as a level floor with tanks at various elevations 
above the floor. Allowing 2§ inches drop between the tanks 
end to end, probably not more than 7 or 8 + anks can be used 
for each circulation system, making blocks of 14-16 tanks, 
occupying a space of 6 to 7 feet by 50 to 65 feet. With the 
Walker system any number of tanks may be placed side by 
side in one row, the circulation being from row to row, which 
are at different levels, and not from tank to tank. Four rows 
are usually arranged in a bay 50 to oo feet wide. 

Cathodes. — The first cathodes* used were of lead-plated 
sheet-iron. In the use of these cathodes it was noticed that a 
preliminary plating of copper prevented corrosion of the iron 
underneath. At Trail a number of tanks were operated for 
some months with ene-eighth inch sheet steel cathodes, but 
the experiment was not regarded as successful. The cathodes 
were provided with grooved wooden strips which fitted on 
their edges, to prevent the growth of lead where it could in- 
terfere with pulling the deposits off. The lack of success 
was on account of lack of sufficient care in the preparation 
and plating of the sheets. Some of the cathodes which were 
carefully plated for depositing starting sheets were not at- 
tacked, but most of the others were. When not well pro- 
tected by copper and lead the iron pitted, and the deposited 
lead was not as hard and solid as when deposited on lead. 

On the other hand the labor was much less with the steel 
cathodes, and there were no short circuits from the anodes 

*U. S. Patent, A. (i. Betts, 679824. August 6, 1901. 



REFINER? CONSTRUCTION, OPERATION, REFINING COSTS. 229 

and cathodes touching. Had they been carefully plated with 
copper before lead-plating them, and re plated if worn out, 
their use would probably not have been abandoned. The 
cost for plant is of course greater with steel cathodes, 
namely about $100 per ton refined per day. Those used at 
Trail were made of tank steel and had to be. selected, as 
some of the steel was too much warped. By stretching, 
perfectly flat sheets could be produced, and this is an actual 
manufactured article I am told, though I have not been able 
to find out where stretched steel sheets are made. A copper 
bolt was riveted and soldered to the cathode lug to take 
the current, while the upper part of the steel cathode was 
painted with P. & B. paint as a protection from acid spatter- 
ing on them. They were greased before receiving the deposit 
so that it could be readily removed. These cathodes may 
be seen in Plate 7. The two round holes were used for 
lifting. 

Usually lead cathodes are used, either of deposited or cast 
sheets. The first cathodes used at Trail were of deposited 
lead, made in four tanks six inches deeper than the others, 
with correspondingly longer anodes and cathodes, the latter 
of copper- and lead-plated steel. These cathodes had been 
carefully prepared, and very good deposits of lead were ob- 
tained, which were stripped off and wrapped by hand around 
the copper cross-bars for the other twenty-four tanks. For 
greasing the steel cathodes a solution of paraffine dissolved 
in benzine was used. It was found necessary to let the ben- 
zine dry off or otherwise the deposited lead stuck fast. 

The rough side of the sheet was put next the cross-bar 
to give better contact,, but the contacts were not as good as 
could be wished. At one time clips were put on the cathode 



230 



LEAD REFINING BY ELECTROLYSIS. 



bars to try to improve the contact, but this was not worth 
the trouble. The clips may be seen in Plate 7. 

A great many plans were suggested for casting thin cathode 











Fig. 42. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 231 

sheets, including rolling, dipping cold iron plates into melted 
Lead, revolving a cooled steel drum in a lead pot, and rolling 
up the resulting lead strip to be afterwards cut into lengths. 
Mr. John F. Miller, of the Canadian Smelting Works, brought 
out the apparatus used at present. Pure lead is kept just 
melted in a small pot and ladled into a pivoted trough at the 
upper end of a sloping iron plate (see Fig. 42). The lead in 
the trough is then tipped on the plate, where most of it solicli- 




Fig. 54. 

fies in a thin even plate, while some is thrown off at the bot- 
tom. These sheets are then thrown on a pile, and later on 
wrapped around the cathode cross-bars by hand. 

Dr. Wm. Valentine has improved this cathode by casting 
two lugs on at the bottom of the plate at the same time the 
plate itself is cast, using suitable molds in connection with 
the plate. His cathode is illustrated in Fig. 43. The cathode 
rod, which is round except where flattened at one end for con- 
tact with the bus bar, is inserted in the holes in the lugs. Fig. 
44 explains the operation of the apparatus used. 



232 



LEAD REFINING BY ELECTYOLYSIS. 



This gives a suspension 
from the center of the cathode 
bar, while by the old method 
the lead is suspended from one 
side, and the rod has to be 
kept from turning when rest- 
ing on the tank and being car- 
ried by the crane, by giving 




Fig. 44. 

the rod a special shape, and 
using a special hook on the 
crane (see Fig. 45). The con- 
tact with the Valentine cathode 
is almost perfect, which is a 
further advantage. 

The lead cathodes are all 
too flimsy and require straight- 
ening before use, and careful 
handling to the tanks, and then 
there are sure to be short cir- 
cuits. In tanks where no short 




REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 233 

circuit actually exists, the uneven spacing of the electrodes 

causes the anodes to dissolve unevenly, which is a bad thing 
for a number of reasons. Methods that will insure an even 
spacing of electrodes and uniform contacts arc worth stri- 
ving for in electrolytic refining. 

Cathode bars. — These are usually of copper. At Trail 
rods \ inch XI inch on edge were found strong enough. The 
two ends were twisted flat and offset one half inch at the same 
time, as shown in the sketch, to prevent the cathode sheet 
from turning the rod. See Fig. 46. These rods weighed about 

r~ > ! <~n j:d] 

i j 

Fig. 46. 

6 lbs. each. As J sq. in. of copper is more than necessary to 
carry 200 amperes, and copper is so much softer than steel, 
a combination rod is cheaper and better. Steel tubing plated 
with copper about iV-inch thick is also in use. The plating 
can be readily done by any electro-chemist or plater of ordinary 
skiU. 

Tank foundations, supports, and arrangement to catch leaks. 
— Brick piers of sectional area corresponding to their height, 
with a concrete base, and glass plate half inch thick on top 
for insulation, make good supports. Concrete is also good. 
A good way of placing the tanks relative to the piers, for 
the cascade arrangement of tanks, is not to have the piers 
under the tanks, but under the aisles between the tanks, while 
the tanks are carried on heavy cross-beams. See Fig. 47. 
By cutting a small notch from the beams near the piers, any 
acid solution is prevented from running down the piers into 
the ground. It is more difficult to collect any leaks on a 



234 



LEAD REFINING BY ELECTROLYSIS. 



vertical support than that which drops free on the sloping 
boards underneath. 

Cleaning tanks. — The usual plan in a copper refinery is 
to have apparatus arranged so that the slime can be sluiced 
out of the tank. In a lead refinery the conditions are different, 



^^^^m 



"-^^^ 




Fig. 47. 

for the lead slime is denser and heavier, and generally only a 
small proportion drops from the anode anyway, and it is often 
removed in a separate cleaning tank. At Trail we tried the 
plan of sluicing slime into a tank car in the cellar, but the appa- 
ratus was not well arranged, and the slime was too thick to 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 235 

run out. For this mot hod of removing slime, which has a gcod 
deal to recommend it, a tank with hopper-shaped bottom 
ought to be used. This is difficult to make of wood, but there 
would be no difficulty in making a concrete tank with such 
a bottom. See Figs. 47 and 48. 

The usual method of emptying a tank is to take out the 
cathodes first and then the anodes. The clear solution is 




Fig. 48. 



next siphoned off into the launder leading to the low-level 
storage tanks, and the slime shoveled into a barrel, while the 
tank may next be cleaned with a sponge. 

On the method of cleaning tanks adopted depends the 
height of the cellar. In one case there must be head room 
in the cellar, the expedient of sluicing the slime all the way 
to a common collecting point requiring too steep a pitch, so 
tank cars must be used which can be run under any tank. 
This means a more expensive plant for excavation and pillars, 
but it has the advantage of diminishing labor cost, and the 
tanks can be washed absolutely clean. When electrolyte is 
added to a dirty tank the slime present is stirred up and set- 



236 LEAD REFINING BY ELECTROLYSIS. 

ties slowly, and a good proportion may be expected to settle 
on the cathodes. The sluicing method will give therefore 
the best results, and the extra first cost is not very great. For 
relatively impure lead, containing say 2% of antimony and 
arsenic, the slime remains so firmly attached to the anodes 
that little or no slime falls into the tanks anyway, and in this 
case the simplest plant will be equally as easy to operate, on 
account of much less frequent cleaning being necessary. 

An arrangement with track underneath for carrying slime 
out is shown in Fig. 47. The only sure way of getting the 
slime to run is to have it drop directly, so the tank car should 
be run directly underneath the tanks. 

Floors of a mixture of asphalt and barite are used and are 
expected to be solution-tight, but it is doubtful if an entirely 
solution-tight floor can be made in this way. There are never- 
theless some excellent and cheap materials available for catch- 
ing the leaks. Ordinary tarred roofing paper, supported on 
boards, is good, and so is roofing paper that has been soaked 
in paraffine. The solution has no effect whatever on the latter. 
The experiment has not been tried, but I feel sure it would 
be successful to cut building paper into squares, soak them 
in paraffine, and lay the squares like shingles on a nicely pre- 
pared sloping surface, either of the ground itself, or the same 
lightly cemented. 

The slime-car arrangement adds to the construction cost, 
as can be readily seen, for extra height and weight of pillars, 
excavation and tracks, by an amount which would probably 
be about $45 per ton per day. Nothing is included in this 
estimate for tank car and haulage apparatus, as this substi- 
tutes for other apparatus in the other plan. Capitalized at 
10% this would be 1.3 cents per ton, while the labor and 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS- 237 

time saved, beside making better lead, would be several times 
that much. 

Contacts.— The plain copper to copper surface has been 
tried and found best. Other methods have been tried and 
given up. Mercury contacts are not good. The mercury- 
disappears rapidly and is probably absorbed by the copper. 
If the copper contacts are sand-papered off, the drop in e.m.f. 
will average about 0.01 volt copper to copper, or copper to 
lead anode. For the anodes, letting the anode lug rest di- 
rectly on the copper bus bar, gives a very good contact. The 
under side of the anode lug must be cast flat in order that 
the anode shall hang straight, and this at the same time makes 
it certain that it will hang straight. 

Circulation of electrolyte. — A heavier solution contin- 
ually falls from the anodes when in action, while a lighter 




Rubber Bow 
to_Pump Tan* 



Fig. 49. 



solution rises at the cathodes. Depositing 35 lbs. of lead 
per hour in a tank causes quick decomposition into a heavy 
layer on the bottom and a light one on top. If the current 
is shut off, and the anodes have a layer of slime attached, 
heavy solution diffuses from the slime for some time after- 
ward. The general method of circulation from tank to tank 
with the cascade arrangement is illustrated in Fig. 49. Rubber 
hose 1} inch internal diameter is fitted in the overflow end of 



238 LEAD REFINING BY ELECTROLYSIS. 

one tank and rests in a notch at the inflow end of the other 
An apron of three-quarter inch wood with a half-inch to one 
inch space between it and the end of the tank, insures that 
only the heavy solution at the bottom of the tank can overflow 
to the top of the next. Some trouble has been experienced 
by the wooden apron shrinking and opening its seams so that 
lighter solution can run through. The use of hard- rubber 
tubes has been attempted in place of the aprons, but I do not 
know whether it is so successful in preventing the agitation or 
suction of slime. My idea in using the aprons originally was 
to have a very slow motion of electrolyte at any one place. 
A hard-rubber tube takes up as much distance from the end 
of the tank as the apron, and the objection to cracks open- 
ing in the apron does not amount to much, for they can be 
easily calked up when the tank is emptied. 

Certain precautions are necessary to keep the tanks from 
overflowing on account of the heavier solution at the bottom. 
If the circulation is uninterrupted the difference in level on 
the two sides of the apron will not be much more than half 
an inch. If the circulation is cut off, even when the current 
has been just cut off too, a sufficiently heavier layer may after- 
ward collect (from the diffusion of the heavy solution in the 
anode slime) sufficient to cause the tanks to overflow when 
circulation is again started, instead of forcing the heavy solu- 
tion behind the apron out at the end. It is very inconven- 
ient to have the solution refuse to pass from tank to tank, 
: and about the only thing to do is to take out some electrodes 
« in each tank, or move them to one end, and stir the tank with 

• a stick. It is also possible to siphon off some of the heavy 

• solution at the bottom and let the tank fill with fresh solu- 
tion. In actual refining the overflowing of the tanks is a 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 239 

rare occurrence, but the causes should be kept in mind so that 
it will be rare. 

The volume of solution passed through a 4000 ampere 
tank can not be so small as to allow of the production of much 
variation in density between top and bottom of the tank, 
nor too great to prevent the slime from settling. Five gal- 
lons per minute is a fair amount. The difference in density 
between the top and bottom is about 3%, and sometimes 
about 5% in practice. 

Several kinds of pumps have been in use, giving varying 
satisfaction. Hard-rubber and bronze centrifugal pumps, 
driven by small electric motors and connected in the circula- 
tion system by good rubber hose, are in use and are very good. 
Hard-rubber plunger-pumps are also satisfactory. Air lifts 
using two automatically operated montejus, as described 
in various books, were tried at Trail first, and found to be 
very poor for the purpose. A wooden plunger-pump was 
then hastily constructed and installed and lasted for a con- 
siderable period with occasional repairs. The pump was a 
long square box of 2-inch planks about 5 or 6 inches inside, 
with a square wood plunger, and leather flaps. Solid rubber 
balls about 2J-inch diameter, on a 1J- to lj-inch hole, made 
excellent valves. Iron was used in the construction of the 
pump, and even in the solution itself it lasted a long time. 
Copper may be used in the construction of the pumps, 
especially when in contact with lead, so a modification of the 
wood plunger-pump, using a copper tube with lead plunger 
and lead valve at the bottom, would be sure to make a good 
pump. Fig. 50. A plunger-pump has the advantage of not 
churning any air in with the solution, and can be expected 
) make somewhat purer lead. It is difficult to see how dis- 



240 



LEAD REFINING BY ELECTROLYSIS. 




Fig. 50. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 241 

solved oxygen in the electrolyte could fail to oxidize some 
slime, thereby dissolving a little antimony which would partly 
go into the cathodes. 

The deposition of the overflow tanks and pumps may be 
varied, some methods being shown in Figs. 51 and 52. For 
filling tanks after cleaning out, the arrangement with a stor- 
age tank at a higher level from which tanks that have been 
emptied may be quickly filled, has some advantages, and is 




Fig. 51. 



Centrifugal 
Pump 



ragal n — »=- »_- «_ „_, 



Fig. 52. 



probably the best. The other way of letting tanks fill by 
putting them in the circulation system again, requiring an 
hour or two to fill a tank, while all below in the same series 
are without circulation, is not to be recommended. 

Considerable storage should be provided for electrolyte 
from tanks which may be emptied for cleaning. It is con- 
venient with the cascade arrangement to cut out two or four 
tanks at once; and to provide for two sets of four tanks out, 
with two to spare, or a storage of say 800 cu. ft., will not be 
found excessive. Two tanks of 400 cu. ft. each, so that one 



242 LEAD REFIX1NG BY ELECTROLYSIS. 

can be removed or repaired, would be good practice for wood 
tanks, and a considerably larger number of sulphur-treated 
concrete tanks of the same size as the electrolytic tanks would 
be right for concrete tanks. These latter could be made ab- 
solutely tight, and could be connected together by siphons 
as far as desired, so as to reduce the number of units to be 
considered to one, with the possibility of always cutting out 
any desired tank for repairs. 

These storage tanks are placed at so low a level that all 
liquids from the lead tanks, whether by leaks or siphons, run 
to them by gravity. 

There is also to be provided another set of 350 cubic feet 
capacity at a level above all the tanks so that it may be dis- 
tributed by hose to any part, with a pump to raise solution 
from the low-level storage set to the high storage tank. 

Electrolyte. — The composition of the electrolyte is treated 
in Chapters I and V. The quantity required, with anodes 
spaced 4} inches center to center, current density 15 amperes 
per sq. ft., is about 120 cu. ft. per ton deposited per day. If 
this contains 200 gr. SiF 6 and 80 gr. lead per litre, its cost is 
about SI. 25 per cu. ft, as follows: 

TABLE 89. 

25. 7 lbs. fluorspar at S14 per ton SO. 180 

34. 7 lbs. 66° H,,S0 4 at S12 per ton . 208 

29 lbs. fine quartz at $10 per ton . 145 

Coal, 15 lbs 0.020 

Labor 0.220 

Repairs . 100 

6 lbs. white lead at 6^ cents per pound 0.375 

$1,248 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 243 

The yield on fluorspar and sulphuric acid taken is as- 
sumed to be only 80% and 92% respectively. The item for 
repairs is quite large, as lead storage tanks and condensers 
do not last very well. The cost of the acid itself without the 
white lead is $0.87 per cubic foot of electrolyte, or 7 cents 
per lb. anhydrous H 2 SiF 6 , the latter item being of interest 
because H 2 SiF 6 only is needed for renewals to keep up the 
strength of the electrolyte. Electrolyte with 160 gr. SiF 6 
and 64 gr. Pb would cost about $1.00 per cu. ft. 

By dissolving lead in the solution electrolytically, instead 
of using white lead, the cost of electrolyte could be reduced 



^ 




Fig. 53. 

by perhaps 10-15 cents per cu. ft. A very simple arrange- 
ment of tanks is required, and the power necessary is small, 
namely, about 25-ampere days per cubic foot. For a 100 
ton per day refinery the hydrofluoric acid would be made 
slowly, requiring say two months or more. The power for 
dissolving the lead electrolytically could be supplied for ex- 
ample by a 12-volt 500-ampere generator, operating 10 cells 
in series, putting the necessary lead into solution as fast as 
the acid was made. See Fig. 53. Current densities of 50 
amperes per sq. ft. would be allowable, and the cells would 



244 LEAD REFINING BY ELECTROLYSIS. 

then be only about 4 ft. square over all. A purer electro- 
lyte to start with may be produced at less cost in this way, 
and it is therefore to be recommended. 

Washing appliances for electrodes. — It has been the cus- 
tom to inspect each finished cathode separately to get off any 
patches of slime, but these patches come from bad work or 
crudely cast anodes and flimsy cathodes touching the anodes. 
This, while excusable in early work, is no longer necessary, 
so that the inspection of individual cathodes and brushing 
where slime is attached is no longer necessary when washing. 
The copper refiners spray a whole tank-load of cathodes at 
once with hot water from a set of perforated pipes, between 
which the crane lowers and raises a tank-load of plates. Lead 
can of course be washed the same way. Dipping cathodes 
into a tank of wash-water does not give so complete a wash 
with a given quantity of water and is more troublesome. 

Anode scrap may be cleaned before removing it from 
the refining-tank, by standing over the tank and passing a 
wiper over the surface of each plate to loosen the slime, pro- 
vided the slime is uniform and not too hard. With drossy 
anodes, hard spots are found in the slime that do not come 
off very easily, and such anodes have been wiped by hand 
in special cleaning-tanks. For slime which remains attached 
securely enough to stand removing with the anode scrap (as 
is usually the case) wiping apparatus, such as shown in Fig. 
54, is recommended. This would not work with slime from 
very hard lead, say 3-10% Sb, as this has to be scraped 
off. 

Regarding the presence of hard particles of slime, this 
results on one side of the anode only where the dross collects 
during cooling. This is obviated by the use of closed molds, 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 245 



and also by casting the lead anodes from the melting-pot at 
a low temperature, leaving the dross undisturbed on the sur- 






1 -f 



1 / 



!_/jl/f I A 



l Ml 



&r~P. 



—7 S ■>■ /■ 




Fig. 54. 



face until most of the lead has been cast, and then raising 
the heat enough to soften the dross so that it may be dipped 



246 LEAD REFINING BY ELECTROLYSIS. 

into a separate lot of anodes, to be refined in a few special 
tanks and afterward cleaned of slime by themselves. 

After removing the slime the anode scrap is sprayed, dried 
and melted, while the wash- water is used in washing slime. 
The wash-water from the cathodes is sometimes used over 
and' over until it nearly reaches the strength of the electro- 
lyte, when it is added to the tanks. For loss of solution in- 
volved, see page 39. 

Slime washing. — The slime, as removed from the anodes, 
contains a large amount of valuable solution which is stronger 
and more neutral than the main body of the electrolyte. Fil- 
tration and washing has been done by suction filters, filter 
presses with iron plates, and decantation. The suction fil- 
ter will probably not come into use any more, although it 
is successful. The slime filters very well in a press, but there 
are difficulties in forcing it into a press, on account of lumps 
of lead that stop up the pipes. Washing by decantation is 
the best in my opinion. To secure the best result, the wash- 
ing should be done on the counter-current principle. One 
washing-tank and several storage tanks for various strengths 
of wash-water comprise the necessary apparatus, with a steam- 
pipe to heat the slime and solution, the latter making it break 
up and wash better. 

Mr. F. C. Ryan, of the United States Metals Refining Co., 
made an experiment as follows, which shows that the heating 
does not decompose any of the fluosilicic acid. 

Equal weights of raw slime were stirred with equal quan- 
tities of hot (180° F.) and cold water, for half an hour, when 
he wash- water was decanted and tested. 

As will be seen there was no material difference in the 
effectiveness of the washing. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 247 

TABLE 90. 



Gold Water 




Hot Water 


Sp.gr. 1.079 = 10.5° B 


Sp. gr. 


1.083 = 11.0° B 


SiF 6 = 2.95% 


SiF 


= 3.0% 


Pb. = 4.76% 


Pb 


= 4.94% 



Experiment indicates that the slime is readily washed, 
and no absorption or retention of stronger solution takes place 
when the slime is stirred up well with w r ater or solution. 

I have worked out an equation from which may be closely 
calculated the effectiveness of washing on the counter-current 
principle with four washings. 

Let a = percentage of acid in last wash- water 
b = " " " " third 

c= " " " " second " 

d= " " " " first 

x= u it ti n solution to be washed from 

slime. 

volume wet slime after settling 
Total volume after adding wash- water* 

The equations are: 



a = 



l-32/+4i/ 2 -2?/ 3 +2/ 4 
b = ^- 

y 

b — a 
c= +a 



d= + b 

y 



a — c 

x = + c 

y 



248 LEAD REFINING BY ELECTROLYSIS. 

If for example one cubic foot of slime is washed with 1J 
cubic feet of water y=$. The acid of the strong solution 
is about 20%, so I have taken x=20. In this case, from the 
above equations: 

a= 1.52% 
b= 3.80% 
c= 7.21% 
d= 12.33% 

This shows a removal of all but 7.5% of the contained 
electrolyte, or say 1.3 lbs. SiF 6 per ton of lead. The amount 
of wash-water to be added to tanks, above the volume of the 
slime taken out, would be only about enough to make up the 
normal evaporation in the tanks, taking y = \. 

If 2/=}, a = 4.00% = 3.2 lbs. SiF 6 per ton lead 

y =i } a = .64% = 0.5 lbs. SiF 6 " " " 

y = i } a = .17%= .14 lbs. SiF 6 ' 

y =^ } a = .06%= .05 lbs. SiF 6 ' 



C (I I ( 



C (I ( i 



Plant for washing is shown in Fig. 55. 

A single plunger-pump can be used for pumping wash- 
water from the storage tanks to the washing tank. The 
storage tanks are at a lower level, so the clear solution may 
be siphoned off directly. The storage tanks are required to 
hold twice as much as the washing tank. 

Conductors. — Rolled copper conductors are used, which 
may be either nearly square in cross-section or flat. The flat 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 249 



]g) Storage for 
wash water 



Siphon 
to empty washer 



Storage for 
wash water 




Fia. 55. 



250 LEAD REFINING BY ELECTROLYSIS. 

bars cool a little more on account of greater surface, increas- 
ing their conductivity slightly. Conductors are either placed 
on top of the tanks or at the side, the best position being on 
top, as a shorter lug may be used on the electrodes. A bar 
1 or 1J inches thick and 4 inches wide is suitable. Any neces- 
sary bends can be put in by heating the bar to a dull red at 
the right place, when the bend can be easily put in. 

Cranes. — The three-motor cranes, though more expensive 
than one-motor cranes, are to be preferred. Those in use 
have about a 50-foot span or more and can carry 10 tons nom- 
inally. Some cranes have one wire hoisting-rope and one 
hoisting-drum, and others have a rigid construction with 
heavy guides at the ends of the electrode racks with a hoist 
at each end; but they are more expensive, although working 
faster than the single-rope type. The cranes carry a sepa- 
rate frame with hooks for lifting electrodes, the point of en- 
gagement for the anodes being underneath the lug just in- 
side the tank, while the cathodes may be lifted at various 
places according to the type of cathode. 

The set of hooks shown in Fig. 56 has its center of gravity 
too near the hook to be satisfactory, but it was made this 
way on account of limited head-room. 

The tank-room floor has a fairly large space, about 15 ft. 
or more in width, at each end for working on electrodes and 
for the industrial railway. A number of racks for holding 
fresh anodes properly spaced for the tanks is convenient. 
These may be brought from the melting-plant best by a crane 
which commands both melting- and depositing-floors, but also 
on the industrial railway. A form of rack to economize space 
is shown in Fig. 57. Before removing the lowest set of anodes, 
the I-beams for the upper set are lifted out of the way. Three 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS 251 






sh> 



» 



M< j *" 



~S~1 
^5 



u 



iir 



•?+>( 



j#^3 



S 



7^ 



-U 



"^ 1 



iJ 






^^ 



^^ 



n 





252 LEAD REFINING BY ELECTROLYSIS. 

sets can be stacked as well as two. The cathodes are dumped 
on flat cars to be taken to the refinery after the supporting 
bars have been pulled out. 

Floors. — The floor planking is not nailed down at places 
near the tanks, to facilitate their removal when cleaning or 
when repairs are necessary. 

Evaporators. — Wood tanks and lead pans, with a steam- 
coil, and also copper pans, have been used for evaporating. 
None of these are perfectly satisfactory. A copper pan was 
used at Trail, but copper was dissolved and the refined lead 
contained copper. It is my opinion, though, that a copper 
pan can be used successfully by properly protecting it from 
dissolving. This can be done by having metallic lead in the 
solution in contact with the pan. Under these circumstances 
no copper could dissolve until there was a difference of 
e.m.f. between the lead and any point in the copper pan of 
about .5 volt. Another method would be to hang a lead pig 
in the evaporator, connecting the pig as anode and the pan* 
as cathode; by passing a small current the pan could be kept 
covered wherever wet by the solution with a little lead, and 
there would be no chance for copper to dissolve on account 
of the considerable difference of dissolving e.m.f. between 
lead and copper. 

Of course the acid water condensing on the upper part 
of the pan could dissolve copper, but there would be no 
trouble in curing this by hanging sheet lead around the 
sides of the pan. 

Wood tanks are not satisfactory for this purpose, while 
lead pans are, though they do not last long. Their life can 
be increased by hanging sheet lead over the sides, or by keep- 
ing a small current passing with a lead pig as anode, as sug- 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 253 

gested for the copper pan, so that the tendency is to thicken 
the pan instead of making it thinner. 

A lead steam-coil is satisfactory as a means of heating the 
solution. The lead dissolves from the coil slowly, but this 
does not make any serious difference. At Trail, when refin- 
ing 20 tons per day, one evaporator 20 inches deep, 8 feet 
wide and 10 feet long, was sufficient to evaporate the wash- 
water from 2 to 10° B. up to 20° B. It should be noted that 
the removal of the slime itself reduces the volume of elec- 
trolyte in the refining tanks, and as this slime is finally re- 
moved wet the volume of the contained wash-water should 
not be lost sight of in calculating the amount of evaporation 
necessary. Some evaporation takes place from the lead re- 
fining-tanks, which I estimate at about 2.2 cu. ft. per ton 
refined per clay. 

The stronger wash-water, if possible, should be added 
back to the cells without being heated, and only the weaker 
solutions evaporated to save losses by volatilization. By wash- 
ing the slime by decantation, and a carefully arranged method 
of washing, I think it would be possible to get along with- 
out any evaporation at all. In fact this was done at Trail 
at first when the acid loss was as follows: 

Aug. 3d— Sept. 16th, 1903, 13.8 lbs. SiF 6 per ton lead 
Sept, 16th— Oct. 6th, 1903, 7.7 " SiF 6 " " " 

Part of this loss was due to absorption and leakage, and 
no adequate means for collecting leaks was provided. 

Regarding the amount of evaporation from the deposit- 
ing tanks, I think it veiy probable that with a well-syste- 
matized washing plan, the evaporation from the tanks will 
take care of all or almost all of the wash-water it is necessary 



254 LEAD REFINING BY ELECTROLYSIS. 

to use. Just how much evaporation takes place I do not be- 
lieve has been determined, because of the difficulty in making 
such a determination in a refinery. We have certain ways 
of getting at this, however. The voltage between electrodes 
in the solution being taken at .3, this is a measure of the elec- 
tric energy expended, which is all absorbed in heating the 
solution, and this serves to maintain the electrolyte at about 
30° C. while the temperature of the room is probably about 
17i° C. A large proportion of the cooling of the electrolyte 
is undoubtedly the result of evaporation. Taking an evapo- 
rative efficiency of only 50%, the volume of water driven 
off per ton lead refined per day would be about 2.2 cu. ft. On 
the assumption that the cooling air (that is the tank-room air 
in this case) is saturated with water, and that it escapes from 
the surface of the liquid three-quarters saturated, Hausbrand's 
tables* give for air temperature 15°, and water temperature 
30°, 65% of the heat absorbed by evaporation, and 35% by 
heating, and for 20° and 30°, 60% by evaporation and 40% 
by heating. My assumption of 50% still allows something 
for heat loss through the sides of the tanks, especially as the 
tank-room air is not saturated with water by any means, 
except occasionally. 

This will easily take care of all the water that need be 
used in washing the slime, and probably of all the wash- 
water needed altogether. 

One cubic foot of ordinary bullion gives about one half 
cubic foot of wet slime after its removal from the 
anode. 



* Hausbrand, " Evaporating, Condensing and Cooling Apparatus/ 
page 327. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 255 



TABLE 91. 



Plant. 


Present 
Capacity. 


Number 

of 
Tanks. 


Electrolyte 

Grams per 

100 cc. 


Electrode 
Separation. 


Current 
Density, 
Amperes 
per Sq.Ft. 


1. Consolidated Mining 
and Smelting Com- 
pany of Canada, Trail, 
B. 0. 


Approxi- 
mately 80 
tons per day 


240 


6-7 Pb 

12-13 SiFe 


4 J ins. 


16 


2. United States Metals 
Refining Company, 
Grasselli, Indiana. 


Approxi- 
mately 85 
tons per day 


176 


7 Pb 

13 + SiFe 


4f ins. 


12-15-1- 


3. Newcastle - on - Tyne 
Plant. Data withheld 
at request of owners. 













Plant. 


Anodes 
Active 
Surface. 


Average 

Volts 
per Tank 


Percentage 
Anode 
Scrap. 


Cathodes . 


Anodes 

per 
Tank. 


Tank 
Arrange- 
ment. 


1. Consolidated Mining 
and Smelting Com- 
pany of Canada, Trail, 
B.C. 


26 XSOh 

ins. 
Weight 
350 lbs. 


.30 


15 or less. 


2 sets for 
each set 
anodes. 
Weight 
150 lbs. 
each. 


20 


Cascade. 


2. United States Metals 
Refining Company, 
Grasselli, Indiana. 


2X3 ft. 
Weight 
400 lbs. 


.38 


25, to be 
reduced to 

15%. 


2 sets. 
Weight 
150 to 
175 lbs. 


28 


Walker 
System. 



Plant. 


Slime Treatment. 


Source 

of 
Power. 


Size of Tanks 
Inside. 


Genera- 
tors 


1. Consolidated Mining 
and Smelting Com- 
pany of Canada.Trail, 
B.C. 


Leaching with sodium 
sulphide solution. 

Melting residue to dore 
after oxidizing sul- 
phides. 


Water. 


84iX 30X44 ins. 


1-3500 

amperes - , 

60-110 

volts. 


2. United States Metals 
Refining Company, 
Grasselli, Indiana. 


Melting to slag, matte 
and dore\ 


Steam. 


132X30X43 ins. 


1-4500 

amperes, 

60 

volts. 



Power plant. — The subject of power plants calls for no 
special remarks here, as more accurate information on that 
subject that I could give can be got elsewhere. In mak- 
ing estimates of refining cost, the power is usually consid- 
ered as being supplied separately, and the item for power 
includes all expenses, interest, and depreciation for the power 
plant. 



256 



LEAD REFINING BY ELECTROLYSIS. 



Slime plant. — All the apparatus mentioned is not appli- 
cable to any one process. 

Drying slime. — This has been done at Trail by filling into 
wheelbarrows and running them into a warm brick oven and 
leaving until dry. On dumping the slime into a brick stall 
it takes fire and roasts itself. It may also be spread on an 
iron floor, or even in a lead pan gently heated from under- 
neath. Apparatus, as shown in Fig. 58, would be an improve- 




Fig. 58. 



ment on the above methods. The heat could be main- 
tained to either dry the slime or roast it, as desired. This 
apparatus is also applicable to roasting with sulphuric 
acid. 

Melting slime. — Magnesia-lined reverbatory furnaces are 
most used for high temperature work in melting and refining 
dore. Most varieties of crucibles, including graphite, are 
rapidly corroded with most slime mixtures. Clay-lined graphite 
crucibles, I understand, are about the best for the purpose. 
For melting slime to matte and slag, or metal, matte and 
slag, iron pots are quite satisfactory, though there is some 
corrosion of the pot by the slag. Pots arranged as shown in 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 257 

Fig. 59, while they have not been practically tried, could not 
very well fail to work, because the metal and matte have very 
little or no action on iron at the moderate temperatures used. 
For melting slime, from which copper, antimony, and ar- 
senic have been removed by wet methods, a silicious slag 




Fig. 59. 



should be produced by reacting on the lead sulphate of the 
slime with silica which is added. This melting can be easily 
done in crucibles, as no furnace refining is required. 

Leaching slime. — The treating of slime with ferric sul- 
phate solution, etc., can be done in lead-lined stir-tanks. 
The solution can be afterward removed by settling and siphon- 
ing off the clear liquid. Washing can be done by decantation, 
or a lead-lined montejus may be used to force the solution 
into a filter-press. At this period of operation, filter-press- 
ing works well. 

Leaching with hydrofluoric acid can be executed in the 
same tank, if washing by decantation has been resorted to, 
or in a smaller tank of the same character if filter cakes are 
being leached. After hydrofluoric acid has been applied in 
excess, the insoluble residue becomes flocculent and easily 
suspended, so that agitation by air would be successful here. 
It is occasionally necessary to add metallic antimony to pre- 
cipitate dissolved silver, which happens mainly if the slime 



258 LEAD REFINING BY ELECTROLYSIS. 

has been air-oxidized. Placing chunks of antimony on the 
bottom of the tank will do, but suspending them with copper 
wires is more convenient. 

Filtration of the antimony fluoride solution is quite simple, 
but the solution has too much corrosive action on metals, ex- 
cept possibly lead, to permit the use of anything but wood 
for filters. A gravity filter, with a cloth supported on per- 
forated lead or grooved wood, is successful though slow. 

The roasting of leady-copper matte, and leaching it with 
iron and copper sulphate solution containing free sulphuric 
acid, is directly analogous to the production of bluestone from 
matte and sulphuric acid. The most successful methods and 
apparatus appear to be those described in "The Mineral 
Industry," Vol. VIII, page 189, and Vol. X, page 231, by 
0. Hofmann. The pulverizing of the raw matte to 50 mesh 
is done by a Krupp ball-mill, the roasting in a two-story Pearce 
furnace; the regrinding is an easier matter, but the method 
is not described. 

The dissolving is done in wood stir-tanks 12 feet in diam- 
eter and 6 feet deep, with a 12 X 12-inch oak paddle for stir- 
ring. A truncated cone in the center on the bottom 5 feet 
3 inches diameter across the top, and 17 inches high, base 
7 feet 6 inches diameter, also of wood, and filled with sand, 
prevents the matte from piling up in the center of the tank. 
After reaction is complete, the mixture runs into air-pressure 
tanks, and is then forced into wood filter-presses, though hard- 
lead pressures ought to be much better. 

The slightly acid solution is neutralized with a little more 
matte in a deep tank with air-blast for agitation, to remove 
iron, arsenic, antimony, etc. In the present case, however, 
the ferrous sulphate present is desired, so sufficient neu- 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 259 

tralization to remove arsenic, antimony, bismuth, and silica 
without oxidation, is only needed, with no oxidation. 

Electrolytic antimony depositing tanks. — Lead-lined 
wooden tanks well painted are in use and answer well, though 
a sulphur-treated concrete tank is probably better. The 
solution should be cooled with a coil of lead pipe, through 
which water is circulated, connection being made thereto 
through a long hose to prevent grounding the circuit. 

The anodes for depositing antimony consist of soft lead 
rods about three-eighths inch in diameter, wrapped in two or 
three thicknesses of muslin. They are merely suspended from 
copper cross-bars at a distance of three inches apart. These 
cross-bars can be covered with lead or rubber. In the latter 
case the rubber is cut away at the places where the lead 
rod comes in contact with the cross-bar. 

The cathodes consist of sheet copper, which have been 
slightly greased to facilitate the removal of the deposited an- 
timony. The cathodes have a certain small disadvantage, 
for if they are sufficiently greased to permit the easy removal 
of the antimony, the antimony is likely to drop off to some 
extent in the tanks, as it curls away from the cathodes, and 
if they are not greased enough the cathodes have to be bent 
and wrinkled to get all the antimony off. It is, however, not 
necessary to get all the antimony off each time, and if pieces 
fall in the tanks they can be readily collected. A current 
density of 15 amperes per square foot, volts about 2.9 to 3.0, 
and a distance from center to center of cathodes of about 4 
inches may be recommended, though this can perhaps be im- 
proved upon. 

Wrapping the anodes should be done with muslin strips 
put on diagonally, with a string or elastic band around each 



260 



LEAD REFINING BY ELECTROLYSIS. 



end to hold it. The anodes should not be allowed to dry with 
acid on, as this rots the cloth; nor should the solutions be too 
strong or contain too much sulphuric acid, for the same rea- 
son. The best way to do is to keep the anodes in use as con- 
tinuously as possible, and if the tank has to be shut down, fill 
it with water. The anode scrap can be thrown into the lead- 
furnace of course, cloth and all. The production of antimony 
is small, so that individual handling of electrodes with a 
block attached to an overhead trolley parallel to the long 
side of the tanks is all that is necessaiy. The form of tank 
is shown in Fig. 60. 




Electrolytic ferric sulphate tanks. — The chemical and 
electrochemical side of ferric-sulphate production has been 
treated elsewhere. In constructing the electrolytic tanks, the 
following diaphragms are available and practical: Perfor- 
ated lead sheets in pairs with asbestos paper or asbestos board 
between; perforated wooden boards with the holes closed 
with asbestos, and hardened asbestos mill-board. All may 
be used interchangeably as diaphragm plates. From -the 
electrical standpoint, the lead diaphragm is the best, on ac- 
count of the low resistance of these diaphragms consequent 



REFINER? CONSTRUCTION, OPERATION, REFINING costs. 261 

on the largo relative area of the holes, while the hardened 
asbestos hoard has the greatest resistance. The resistances 

have not been accurately measured, but Table 92 (page 2G2), 
from tanks of various sizes in operation, is of interest. 

Of the above tanks all but Xo. 2 gave high current effi- 
ciency; No. 2 had internal leaks and gave a low efficiency. 

Fig. 61 illustrates a tank for 3500 amperes. The cathode 
bus bar runs lengthwise in the center of the tank, and a cath- 




Fig. 61. 



ode is placed on each side in each compartment. The anodes 
are of Acheson graphite one inch in diameter, spaced 1J inches 
centers. The anodes are inserted in the channel irons, and 
cast in lead, which makes a good contact. The sheet-copper 
cathodes, cross bars, and lead lining of the tank call for no 
special remark. The circulation of the anolyte, which is quite 



262 LEAD REF.INING BY ELECTYOLYSIS. 

TABLE 92. 



Diaphragm. 


Date. 


Electrode 
Separation. 


Diaphragm. 


Holes . 


Holes 
Center to 
Center. 


1. Wood and 

Asbestos. 


Sept. 
1903 


3" centers. 


|" wood, bored, and holes 
filled with asbestos. 


h" 




2. Wood and 

Asbestos. 


1905- 
1906 


3^" centers. 


f" wood, bored, and holes 
filled with asbestos. 


i" 


H" 


3. Lead and 

Asbestos. 


June. 
1906 


4i 


2\ lb. lead sheets, asbestos 
paper between. 


i" 


i" 
average. 


4. Sulphurized 
Asbestos. 


1906 


3f 


\" asbestos, with absorbed 
sulphur. 






5. Sulphurized 
Asbestos. 


1906 


3f 


\" asbestos, with absorbed 
sulphur. 







Diaphragm. 



Date. 



Compartment s . 



No 
Com- 
part- 
ment. 



Active 

Cathode 

Area. 



Solution. 



Temper- 
ature. 



1. Wood and 
Asbestos. 



Sept. 
1903. 



18"X13"X3' 



11 



13.8 sq.ft. 



3% H2SO4 
4% Fe" + Fe' 
3% Cu" 



28° C 



40° C. 
41° C. 



Wood and 
Asbestos. 



1905- 
1906. 



33" X5 ft. 



23 



260 sq. ft. 



3.4% H2SO4 
4 % Fe" 
3 % Cu 



40° C. 
50° C. 



3. Lead and 
Asbestos. 



June, 
1906. 



3f"X29i"X12V 



3.33 sq.ft. 



5.5 % H2SO4 
3.5 % Fe 
3.75% Cu 



62° C. 



4. Sulphurized 
Asbestos. 



1906. 



33"X23' 



24 sq. ft. 



4-5% H2SO4 
4% Fe 
3% Cu 



40-50° C. 



5. Sulphurized 
Asbestos. 



1906. 33"X23' 



24 sq. ft. 



4-5% Fe' 



44° C. 



Diaphragm. 


Date. 


Am- 
peres. 


Volts. 


Amperes 
Sq. ft. 


Kind of 

Carbon . 


Notes. 


1. Wood and 
Asbestos. 


Sept. 
1903. 


115 

100 
90 


2.25 

1.60 
1.50 


8 3 Amor- 
phous. 
7.25 ! Carbon. 
6.5 


Polarized stationary anodes. 

Clean anodes, moving, no 
polarization. 


2. Wood and 

Asbestos. 


1905- 
1906 


2000 


2.0 


7.7 


Graphite. 


No polarization. Internal 
leaks. Poor efficiency. 


3. Lead and 
Asbestos. 


June, 
1906. 


8. 
33. 


1. 
1.7 


10.0 


Graphite. 


Normal conditions of cur- 
rent and voltage. 


4. Sulphurized 
Asbestos. 


1906. 


200- 
240 


2.4? 


8.3-10 


Graphite. 


Copper-slime treatment. 


5. Sulphurized 
Asbestos. 


1906. 


140 


1.6 


5-8 


Graphite. 


Lead-slime treatment. Sil- 
ica in solution and anodes 
afterward polarized. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 263 

vigorous between the various anode compartments, and of 
the catholyte which circulates freely around all cathodes, 
is maintained by two separate air-lifts, one for catholyte 
and one for anolyte. The diagrams, Fig. 62, will explain the 
circulation. 

The tank is operated continuously, and the anolyte and 
catholyte do not change in composition, the maintenance 
at a practically constant composition being assured by the 
continual inflow of fresh solution. 

The inflowing solution contains about 3% of copper and 
5% of ferrous iron, beside 2 to 5% of sulphuric acid. The 



■K 



Air Lift 



>, 



+-€ 



>+ 



4-€ 



ANOLYTE CIRCULATION 



>' 





Catholyte 








Anolyte 






— 




*S 




Anolyte 







1 t 



Anolyte 



Catholyte 



^ (*=±^ 



CATHOLYTE CIRCULATION 



Fig. 62. 



catholyte contains about j% of copper and the same amount 
of ferrous iron and acid. 

The level of the catholyte is higher than the anolyte by a 
half inch or so, depending on the diaphragm, and the result of 
constant feed of fresh solution is that catholyte continuously 
flows in a small stream or percolates to the anolyte, which 
assays about the same in free acid and copper as the catho- 
lyte, but contains only 0.8-1.0% of ferrous iron, the rest being 
ferric iron. The result of continuous feed is of course con- 
tinuous overflow of finished ferric sulphate solution, through 
a run-off pipe provided therefor. 



264 



LEAD REFINING BY ELECTROLYSIS. 




Fig. 63. 



In putting the tank together, the main point is not to have 
any internal leaks from catholyte to anolyte and vice versa. 
There is no difficulty about this if the following method 
is adopted. In the first place the dia- 
phragms, if of wood, are of separate 
boards, which should be tonguecl and 
grooved. These boards fit between two 
frames, one on each side. A small piece 
of round asbestos packing should be 
tacked on the distance frames, with small 
brads, before placing the diaphragms. 
When the whole is driven together by 
the end wedges, this makes a good 
joint, See Fig. 63. 

The lift required to circulate the solution 
is only about 3 inches at most, and the 
height can be easily got with a 3-foot depth ' 
of solution in the air and solution pipe. 

The siphons for feeding anolyte to and 
from the various compartments are rather 
hard to manage and keep working unless 
provided with a small pipes at the top for 
drawing air out, as in Fig. 64. The anode 
connection may be made by a copper bar 
lying on the center of the anode frame, di- 
rectly over the cathode bus bar, with large 
wires attached over each channel-iron and 
with the other end buried in the lead- 
A flexible connection is required of course 
from the anode bus bar to the outside source of current. 
The anolyte is a little heavier than the catholyte, and by 




Fig. 64. 



REFINERY CONSTRICTION, OPERATION, REFINING COSTS. 265 

supplying the necessary heat to the anolytc, this may be cor- 
rected a little by its greater heat expansion, thus diminish- 
ing the tendency for mixing, existing especially in deep tanks. 

A lead pipe, heated by steam, lying in one of the anolyte 
troughs at the side of the tank does the heating well. A piece 
of 1$ pipe, 6 ft. long, is enough for a large tank. 

Fig. 65 shows a smaller tank for 250 amperes. The 
hardened asbestos diaphragms are 38 by 25 inches. By 
increasing these to 40 by 40 inches, and putting more com- 
partments into the tank, it could be easily extended to take 
2000 or 3000 amperes. The diaphragms are made as follows: 
Powdered sulphur is sifted evenly over the surface of f-inch. 
asbestos mill-board in the amount required, namely, about 
J lb. of sulphur per square foot for a sheet of this thickness. 
The sheet is then slid into an oven heated by an oil bath to a 
temperature of 120-140° C. for one or two hours, or until all the 
sulphur has melted and soaked in. The sheet is taken out, 
cooled, and an equal amount of sulphur put on the other side 
and heated again. The board is cooled on a perfectly flat 
floor, and makes a hard, slightly elastic and waterproof 
product. 

Before putting the diaphragms into the electrolytic tanks, 
they ought to be soaked about two weeks in very dilute sul- 
phuric acid, as some expansion takes place at first. Other- 
wise the sheet will warp in the tank. The method of assem- 
bling and packing the joints is the same as for the tank just 
described. This construction and diaphragm gives a tank 
which is so tight internally that it allows no mixing of anolyte 
and catholyte, even if there is a considerable difference in 
level between the two, and it is necessary to use a small siphon 
to keep catholyte always flowing to the anolyte. 



266 



LEAD REFINING BY ELECTROLYSIS. 




Storage capacity should be provided at a higher level to 
hold sufficient solution to feed the tanks for 36 to 48 hours, 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 267 

and at a lower Level to receive the overflow for the same time. 
The iron in the solution should not be less than 5% and could 
very likely be 6 or 7%, though this has not been attempted 
yet. The ferrous iron in the overflow should be 0.8 to 1.0%, 
the remainder being ferric iron. With 50 gr. total iron per 
litre 1000 amperes produces about 45 cubic feet per day per 
tank containing 10 gr. ferrous iron per 1000 cc. 

Electrolytic tanks, depositing copper from sulphate solu- 
tion with insoluble lead anodes and no diaphragm, are too 
simple to call for much remark. In the execution of the fer- 
ric sulphate process, after neutralizing with matte, electro- 
lysis of the solution with a lead anode, until a few per cent 
of free sulphuric acid is present, could be practiced, and 
would be more economical unless sulphuric acid was very 
cheap. 

Refinery operation and costs.— 1 The tank-room opera- 
tions can be arranged so that little labor is required on Sun- 
day or at night. The daily operation includes charging and 
emptying a certain number of tanks and drawing cathodes 
and replacing them in an equal number of other tanks, the 
practice being to make two sets of cathodes from each set 
of anodes. The operation of changing cathodes is simple, 
but requires care to keep the old cathodes from wiping slime 
from the anodes as they are pulled out, and the new cathodes 
on account of their usual flimsy character have to be handled 
delicately. In view of the greater ease of handling and supe- 
rior electrical and chemical results, I believe a steel cathode, 
as described on page 228, with wooden strips 1 inch square 
having a groove on one side, slipped over the edges, ought 
to be used, though they are not now\ These strips may be seen 
in the photograph, Plate 7. 



268 LEAD REFINING BY ELECTROLYSIS. 

Grooves around the plates will take the place of wood 
strips.* 

It is necessary to inspect the tanks with a voltmeter to 
detect short circuits, and any short-circuited plates are taken 
out and straightened. 

At Trail, B. C, after removing the cathodes, the anodes 
are taken out, a tank-load at a time, and the anode slime is 
removed in separate tanks, by wiping the scrap with rubbers 
and pouring water on afterward to clean off the muddy solu- 
tion. Another way is to hang the whole tank-load in a special 
tank to receive the slime, and reach down between the plates 
with wipers to loosen the slime, after which a spray is turned 
on the plates. This is quite readily done, but apparatus shown 
in Fig. 54, is believed to be better yet, although not yet in 
use. This apparatus will clean a whole tank-load at once, after 
which they may be sprayed with a set of spray-pipes. 

After the anodes have been taken from the tank, the cir- 
culation of electrolyte is shut off, by connecting the overflow of 
the tank next above and the feed of the one next below with 
a hose. The clean solution is then siphoned from the tank into 
a launder beneath the tanks which carries the solution to the 
storage tank. The workman, with rubber boots on, next gets 
into the tank and shovels the slime into a barrel. This is 
more troublesome and less satisfactory than sluicing the slime 
into a car beneath the tank, because the tank should be abso- 
lutely clean when solution is next admitted. Otherwise the 
remaining slime is stirred up, which is a bad thing for the 
cathodes. It should be remarked, however, that it is not 
necessary to clean out a tank at the end of each run as a gen- 

* U. S. Patent, Elliott and Kishner, 683283. October 12, 1901. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 269 

eral thing, for most or all the slime comes out on the anode 
scrap, unless the anode bullion is unusually pure. Usually 
the anode hardly changes its appearance during the whole 
depositing operation. To fill the cleaned tank, it is only neces- 
ary to run a hose to it from the high-level storage tank and 
start the solution by opening a valve or by siphoning. The 
temporary hose to carry the circulating solution around that 
tank is then disconnected and taken away. 

The cathodes are best washed by spraying with warm or 
hot water, though they have also been washed by dipping 
into a tank containing wash- water. If the wash- water is 
used over and over, until it reaches the strength of the tank 
solution, there is a loss of solution of course, as it will not 
all drain off. The amount of solution that it takes to wet 
the cathodes varies of course with the cathode thickness and 
roughness. With the samples shown in Plates 2 and 3, pages 
38 and 39, the amount required to wet them, and corre- 
sponding acid loss, with the weight of cathodes per square 
foot, is given in Table 93. 

TABLE 93. 



No. in 
Photo- 
graph. 


Weight per 
Square Foot. 


Solution 

on 
Cathode. 


Acid Loss 
per Ton Lead. 


Actual Loss. 


Remark. 


2 
4 
5 
6 


28.8 lbs. 
22 
16 
1.1 " 


•5 % 

• 39% 

• 36% 

.22% 


1.66 lbs. SiF 6 

1.33 " SiF 6 

1.2 " SiF 6 

.76 " SiF 6 


.83 lbs. SiF 6 
.67 " SiF 6 
.60 " SiF 6 
.38 " SiF 6 


Average cathode. 
Average cathode. 
Unusual cathode- 
Unusual cathode- 
Not well wetted. 



The actual loss, if the cathodes are first washed with wash- 
water, and this is used over and over until the same strength 
as electrolyte, would be one-half the loss if the cathodes are 



270 LEAD REFINING BY ELECTROLYSIS. 

merely drained. At 7J cents per pound of SiF 6 the maximum 
saving is so small, that a more systematic method of washing 
would not be apt to pay. The amount of wash-water to be 
returned to the tank by this method equals the amount taken 
out on the cathodes, so no evaporation is required. 

The surface of the anode scrap is only from one-third to 
one-half that of the two crops of cathodes for each anode, 
and the anode scrap is smoother too, so that the acid loss from 
solution and wash-water required to wet anode scrap, is about 
30 to 40% of that on the cathodes. 

The acid loss in fairly well-washed slime depends some- 
what on the amount of slime For an average grade of bul- 
lion containing 96-97% of impurity, the loss in fairly washed 
slime will not exceed 2 lbs. SiF 6 per ton of lead, so that the 
total losses on and in material removed will not exceed about 
3.1 lbs., if it is that high. When it comes to evaporation, which 
is, however, not absolutely necessary, there is a chance of boil- 
ing off acid. 

The anode scrap and cathodes are usually carried by the 
crane directly to the melting-pots and dumped in, the 
cathode cross-bars of course being first pulled out. 

For disconnecting tanks from the electric circuit while 
cleaning them, a small copper block and a clamp is all that 
is necessary. For disconnecting one tank it is usual to place 
copper rods across from side to side, resting on the conductors 
on each side of the tank. Sometimes the cathode supporting- 
"foars can be used, but usually they are too short to reach, and 
a few bars of special length are necessary. 

The tank inspector has a voltmeter supported by straps 
around the neck and shoulders so that it lies open in front 
of him. The leads are connected to a pair of small ice-picks. 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 271 

With one in each hand, the voltage of all the cathodes in a 
tank can be quickly read. Any short circuits may be indicated 
by chalk marks, to be fixed by moving the electrodes slightly,, 
or if necessary, by taking the cathodes out and straightening 
them. 

Slime is variously washed by decantation, and by filter- 
pressing. The results obtained by washing by decantation, 
are mentioned on page 247. In washing anode slime by decan- 
tation, if hot wash-water is used, the slime breaks up better 
and is more rapidly mixed with the wash-water. 

Elevating slime either to washing-tanks, or to a filter-press, 
should not be attempted by a montejus, this having been a 
partial failure several times. It ought to be mechanically ele- 
vated in tanks, or driven through a good-sized iron pipe with 
a pump. The former of these two methods was in use at Trail 
at first, and is sure, though clumsy. 

The various wash- waters from cathodes and anode scrap 
should be filtered and run to a storage tank, and then evaporated 
if necessary. The strong wash-water from the slime can go 
directly to the electrolyte storage tanks. 

Making cathodes, as invented by Dr Wm. Valentine (see 
page 231), requires one man, who makes and hangs at least 
10 sheets an hour. One man can make 400 sheets of the 
kind used at Trail, or enough for 30 tons of lead, in a day, and 
in eight hours two men can hang and straighten the same 
number of sheets, so that the cost for sheets, with labor at 
$2.00, is about 20 cents per ton. The cost for Valentine 
cathodes is then a little higher, but they have certain advan- 
tages over the old style, in giving better and more uniform 
contacts, and the lugs being made thicker than the plate itself, 
reduces very much the liability of the cathodes being cut 



272 LEAD REFINING BY ELECTROLYSIS. 

through by the electrolyte at the surface, and dropping in 
the tanks. To prevent this with the old style of cathodes, a 
streak of asphaltum paint was put on where the surface of 
the solution comes. There is no doubt but what the molds 
for making Valentine cathodes can be improved so as to save 
considerable labor. 

The labor cost for operating tanks, that is, charging and 
drawing and washing and cleaning electrodes, cleaning tanks 
(on the supposition that the slime is sluiced out into a car 
underneath), inspecting tanks and fixing short circuits, hand- 
ling anode scrap and weighing, may be taken in detail as fol- 
lows, for a production of 100 tons lead per day: 

TABLE 94. 

Charging tanks 4 men 8 cents per ton. 

Emptying tanks 4 " 8 " 

Cleaning tanks 4 " 8 " 

Inspecting tanks 9 ' l , 3 shifts, 18 ' ' 

Weighing and tramming 3 " 6 " 

Cleaning and handling scrap 4 ' ' 8 " 

Repairs 2 " 6 " 

Making and straightening sheets 10 ' ' 20 " 

Other operations 2 " 4 " 

Total tank-room labor 42 men 86 cents per ton. 

By the use of steel cathodes I believe this cost can be 
reduced to about 60 cents per ton, by removing the necessity 
of making starting sheets and of much inspecting, beside im- 
proving results. 

The labor cost of loading pig lead and unloading bullion, 
sampling, weighing and tramming to and from melting plant, 
would be about 19 cents per ton refined. 

In the melting plant, about 100 lbs. of coal or less is used 
per ton of lead refined. The labor cost charging the kettles 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 273 

and molding anodes and lead and stacking the anodes, is about 
as follows: 

TABLE 95. 

Charging furnaces with lead, and skimming 10 cents per ton refined. 

Molding lead, including firing 9 " " " 

Molding anodes and stacking, including firing 13 " " " " 

Repairs, including new pots 6 ' ' " " " 

Coal 10 " " " 

48 cents per ton refined- 



It might be of interest to state the approximate labor cost 
when handling electrodes singly with an overhead trolley, 
hoisting being, done by a chain block, on a scale of 10 tons 
per day. 



TABLE 96. 



Unloading anodes from cars 

Tramming to tank-room 

Straightening and charging anodes. 

Making starting sheets 

Inspecting and night man 

Charging cathodes 

Drawing and washing cathodes. . . . 
Drawing and cleaning anode scrap. 

Molding 

Loading on cars 

Unclassified 



6 cents per ton refined 

4 
30 
25 
45 

8 
12 
40 
25 
10 
40 



Total tank-room labor cost $2 . 45 cents per ton refined. 



The anodes are supposed to come to the refinery already 
cast, and merely need straightening. The figures are per ton 
refined lead produced, assuming a wage of 20 cents an hour. 
The labor could be reduced considerably. 



274 LEAD REFINING BY ELECTROLYSIS. 

Comparative Costs of Refining by the Parkes and Betts 

Processes. 

Parkes Process. — Assuming that all approved labor-saving 
machinery is used, that the bullion contains .7% Sb, .8% Cu, 
and 75 ozs. silver with a little gold, coal at $2.50 and coke at 
$5.00, zinc at 6 cents, and a production of 100 tons per day 
average wages $2.00 per day: 

TABLE 97. 

400 lbs. coal per ton bullion received at works SO . 50 

65 lbs. coke for reducing hard lead, retorting, etc 0. 16 

Zino,16 lbs 0.96 

Repairs and supplies 0.25 

Parting and refining silver and gold 0. 19 

Fluxes 0. 11 

Labor, softening and desilverizing . 23 

Labor, retorting . 07 

Labor, cupelling . 05 

Labor, power plant 0.14 

Labor, working by-products . 34 

Foremen and general labor . 40 

Mechanics and helpers 0.18 

SI. 41 

Labor, except parting plant SI . 41 

Refining charge on 12 lbs. copper . 09 

$3.67 

No published detailed costs of refining as it is done at 
present, exist as far as I know. The above are compiled from 
various sources of published and private information.* 

The above assumptions may be criticised on the ground of 
too high a percentage of copper in the bullion. With .2-3% 
of copper, the costs would be about 15 cents less. 



* I am much indebted to Mr. Ernst F. Eurich for figures which have 
been largely used in compiling the above statement. 



REFINERY CONSTRUCTION, OPERATION, REFININC COSTS. 275 

Betts Process costs on same bullion, with the same assumed 
cost of coal and labor: 

TABLE 98. 

Power 7.6 H.P. days total at $50 per E.H.P. year $1 .06 

Tank-room, platform, and repair labor 0.86 

Melting lead, labor, supplies, repairs 0.38 

Coal for melting lead 0.13 

Chemicals, 6 lbs. SiF 6 , at 6 cents $0 . 36 

! " glue 07 0.43 

Slime treatment, except power and assaying, including parting.... 0.96 

$3.82 
Credit about 20 lbs. electrolytic copper recovered from matte at 3 cents . 60 

Net cost $3 . 22 



For a complete comparison of the two processes it is neces- 
sary to take into account the metal losses, interest on plant, 
and general expenses. The lead loss in the electrolytic process 
is practically none, as even the lead in the slime is returned 
to the lead blast-furnace. Five pounds lead per ton is an 
outside estimate of loss. The zinc process will lose about 1% 
of the actual lead present. The antimony loss is respectively 
10% or less and 40%, while the electrolytic antimony from 
the electrolytic process is also usually more valuable than 
antimony in hard lead. The silver loss should be calculated 
on actual contents, which is 1|% greater (about) than that 
shown by commercial fire-assay. There is no opportunity for 
appreciable silver loss in the electrolytic process, while with the 
zinc process the loss ascertained from various sources of infor- 
mation may be taken at 1% as an average for good work. The 
same figure can be certainly surpassed by the electrolytic 
process, but lacking clean-up figures from lead refineries, I 
assume the same figure for silver loss for the electrolytic process 
as for the zinc process. 



276 



LEAD REFINING BY ELECTROLYSIS. 



TABLE 99. 



Net working cost 

Interest on plant at 10%* 

Lead loss, at 5 cents per lb 

Antimony loss, at 15 cents per lb 

Interest on metal in process, at $150 per ton, 

at 6% f 

Interest on by-products and dore 

Superintendence and assaying 

Silver loss, actual ! 




* Interest on power plant not included, as this is figured as part of power 



cost. 



Other items for expressage, management, taxes, insurance, 
etc., I assume to be practically the same for each process. 
The above estimate applies to the purer grades of bullion, free 
from bismuth. Copper does not average as high as .8% in 
many cases, but that is usually the result of skimming the 
bullion at the smelter, the dross going back to the lead fur- 
naces and yielding copper-lead matte; but the end-result is 
the same whether the copper dross is skimmed by the smelter 
or refiner, for the metallurgical process is the same in each 
case. 

If bullion with more impurity is under consideration, the 
relative advantage of the electrolytic process is greater. For 
example, if the bullion contains say .05% of bismuth, not a 
large or unusual amount, the electrolytic process produces 
corroding lead, while the zinc process does not, making a fur- 
ther difference in this country, apart from the value of the 
bismuth saved, of $2.00 to $2.50 per ton. With higher anti- 
mony the advantage of the electrolytic process again increases, 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 277 

the amount being easily figured from the known difference in 
antimony loss, 10% and 40%, and the value of pure lead 
and antimony, as against that of the same combined in 
hard lead. 

The present aim of lead smelters is to exclude from the 
furnaces ores containing antimony and especially bismuth in 
order to produce as pure bullion as possible. If the lead is 
refined electrolytically, these metals become a source of profit, 
and the way is opened for the utilization of low-grade bismuth 
ores particularly. 

The following table shows the cost of the two processes 
under the head of labor, coal, chemicals and zinc. 

TABLE 100. 





Parkes 
Process. 


Betts Process. 




Steam Power. 


Gas Power. 


Labor 

Fuel for all purposes at $2.50. . . . 

Chemicals 

Zinc 


$1.41 
0.75 
0.05 
0.96 


$1.89 
0.58 
0.66 


$1.89 
0.35 
0.66 







This shows labor to be less with the Parkes process, while 
fuel and materials are less for the electrolytic process. I have 
assumed that the coal is of good quality and 2 lbs. are required 
to generate 1 E.H.P. hour with steam, and 1 lb. with gas 
engines. The fuel for the Parkes process -includes the fuel for 
treating by-products up to and including fuel used in refining 
copper. 

The following estimates of cost of a refinery to treat 50 
tons of bullion per day, with a maximum capacity of 60 tons, 



278 LEAD REFINING BY ELECTROLYSIS. 

will serve as a basis for other calculations under special con- 
ditions. 

The cost of construction is greatly different with different 
arrangements of plant, cascade system, Walker system or 
series system, and slime plant. The following figures apply 
to the Walker system. The most economical arrangement of 
tank plant and melting plant I believe to be is to have the 
two parts under the same roof, in a long building and com- 
manded by the same cranes. The anodes can then be taken 
directly from the casting floor to the tanks without rehand- 
ling, and the cathodes, after spraying, can be dumped directly 
from the tanks, in or near the melting furnace, depending on 
the kind of furnace. To save in the number of trips required 
of the crane, which would have to be operated steadily to load 
and unload as much metal as 60 tons per day from melting 
floor to tanks and back, the tanks would be made of the largest 
practicable size, to take 6500 amperes, for 60 tons production. 
There would be two cranes on the single runway, in a building 
55X250 feet, of which 100 feet in length would be occupied 
by the tanks. The melting room need not necessarily be 
limited to the same width as the tank floor. 

The cathodes are assumed to be of lead-plated steel, which 
will save enough in a year in operating cost to pay for them- 
selves. The current density to be 15 amperes per square 
foot for 50 tons per clay, and 18 amperes for 60 tons per day. 
The current efficiency is assumed to be 90%, and will probably 
average 95% with steel cathodes. 

The power plant would be required to deliver a maximum 
of 6500 amperes and 43 volts ( = 280 K.W.) to the depositing 
tanks. If only 50 tons per day were produced 197 K.W. are 
required, and for 40 tons per day 128 K.W. The power plant, 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 279 

if in a single unit, should be capable of operating efficiently 
from less than half its maximum capacity, all the way up. 
The power for treating slime and general power and lighting 
will be included in the estimate later. 



Power plant for lead depositing at $135 per K.W $38,000 

104 depositing tanks 3 feet wide, 3 feet 10 inches deep, 8 feet 

6 inches long inside 5,200 

200 feet | steel rods $4.00 

Labor on concrete 10 . 00 

Molds expenses 2 . 00 

3 barrels cement 4 . 50 

22 cubic feet sand 1 . 00 

43 cubic feet rock 2 . 40 

375 lbs. sulphur 4 . 70 

Fuel .50 

Paint 1.00 

Concrete piers and beams 12 . 00 

Labor coating tanks 3 . 00 

$45.10 
Wood tanks are more expensive. 

Lumber, 650 ft. yellow pine at $35.00 $27.50 

Labor, 50 hours 15 . 00 

Iron, 200 lbs 8.00 

Paint 1.50 

Piers and timber supports 10 . 00 

$62.00 



Electrolyte, about 7000 cubic feet 7,000 

70 tons fluorspar at $14.50 $980 

80 tons sulphuric acid at $15 1,200 

18 tons fine quartz at $20 360 

20 tons white lead at $120 2,400 

Labor 1,250 

Repairs 500 

Coal 200 

$6,890 



280 LEAD REFINING BY ELECTROLYSIS. 

For grading and preparing solution-tight floor under tanks, on a 

level site, about 1,000 

Tank part of building 55 X 140 ft. at $1.25 per square foot for walls 

and roof 9,600 

2 electric cranes installed 12,000 

Copper for bus bars at 25 cents per lb 1,250 

Concrete electrolytic storage tanks 500 

2400 steel cathodes \" thick at 3 cents 5,400 

Labor and material for same 1,200 

Pumps, hose, cleaning tanks, electrode racks, starting-sheet appara- 
ratus, evaporator, slime-washing tanks, lights, water connec- 
tions, tracks, cars, sulphur tank, total 5,000 

Royalty for use of Walker system 

Hydrofluoric-acid plant. This is quite cheap to instal, and may 

be expected to cost $1,500 or less for a good-sized plant 1,500 

Total for tank plant, exclusive of royalty $49,650 



The cascade arrangement would cost more, about as 
follows : 

For more building $3,000 

For more copper at 25 cents 2,275 

For power plant to supply power lost in conductors 1,690 

$6,965 
Melting plant costs: 

3 60-ton kettles complete with stack 3,110 

Cast iron, at 3 cents $1,300 

55,000 red brick, at $10 550 

12,000 fire-brick, at $30 360 

Mason's labor 700 

Supplies 100 

Reinforcing iron 100 

$3,110 

Building, about 6,000 sq. ft 8,000 

Molds, open 275 

Tracks, cars and hoists, crane runway, etc 3,000 

$14,385 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 281 

Slime plant, to treat daily GOO lbs. copper, 1,200 lbs. antimony, 
400 lbs. arsenic, 3,750 ozs. silver and gold, and 250 lbs. lead, in 
slime, beside 2 tons of copper-lead matte: 

Slime-dissolving tanks for ferric solution, total capacity 1,000 cu. ft. 1,200 

Antimony-dissolving tanks, 150 cu ft 300 

Lead filter-press, with montejus, for slime 600 

Storage tanks, lead-lined, for sulphate and fluoride solutions, 3,500 

cu. ft. capacity 1,500 

18 2,000-amp. copper-iron electrolytic tanks 6,300 

10 2,000-amp. antimony tanks with cathodes 1,250 

Crucible melting furnaces for antimony, gold, silver, dore, with 

molds 1,000 

Parting plant 600 

Building, about 5,000 sq. ft., at $1.50 per sq. ft 7,500 

Dissolving tank for matte 600 

Filter press and montejus for matte 600 

Roasting furnace 1,000 

Accessory apparatus 2,000 

Filters for antimony solution 100 

Mill for grinding matte 500 

$25,050 

Power plant for treating slime, capacity 120 K.W., at $135. . 16,200 
For general purposes, 30 H.P., at $135 4,050 

Total costs of refinery, maximum capacity 60 tons per day, would 
then be: 

Power plant $58,200 

Tank plant 49,150 

Melting plant 14,385 

Slime plant 25,050 

$146,785 
Engineering expenses, railroad facilities, land, contingencies not included. 

A series plant for a maximum production of 60 tons per 
day, provided with a plant to treat slime by the roasting-with- 
sulphuric-acid process, would work out about as follows: 

Maximum current density 16 amperes per square foot. 
Thickness of electrodes to be J inch, and spaced 1£ inches apart. 



282 LEAD REFINING BY ELECTROLYSIS. 

Volts per plate .22, efficiency 90%. Anodes 3 feet square. 
Tanks 4J feet deep, 3 feet 2 inches wide, and 8 feet 4 inches 
long, taking 56 plates and producing with 144 amperes, 1,450 
lbs. of lead per day, or refining about 1,490 lbs. of bullion. 
88 tanks arranged in 11 sets of 8 tanks each, 10 sets always 
in use, absorbing altogether 1,440 amperes and 100 volts. 

Lead-depositing power plant 145 K.W., at $135 $19,600 

88 tanks of concrete, at $55 $4,840 

Electrolyte, about 8,000 cu. ft 8,000 

2 electric cranes 12,000 

Copper conductors, 2,500 lbs 625 

Preparing floor under tanks 1,000 

Building, 55-125 feet at $1.25 per sq. ft 8,500 

Hydrofluoric-acid plant 1,500 

Electrolyte storage tanks 500 

Accessories 5,000 

Tank room and equipment $41,965 

Melting plant, using rolls to roll anodes or closed molds $22,000 

Slime plant, using roasting with sulphuric-acid process: 

Mixer for slime and H 2 S0 4 $250 

Flat cars and oven for drying slime 2,700 

Electrolytic copper tanks, 15 for 700 amp 800 

Electrolytic antimony tanks, 23 for 700 amp 1,500 

Dissolving tanks, with stirring-gear 800 

Filter press and montejus 600 

Storage tanks 1,200 

Evaporators f cfr H^S0 4 . . 200 

Crucible melting furnaces 1,000 

Parting plant 600 

Building about 3,000 sq. ft., at $1.50 4,500 

Accessories 2,000 



Total $16,150 

Power for slime treatment, 75 K.W. 
' ' and lights, 30 K.W. 

105 K.W. at $135 $14,200 



REFINERY CONSTRUCTION, OPERATION, REFINING COSTS. 283 

For a comparison between the two methods of installation 
we have for plants with 60 tons maximum capacity: 

TABLE 101. 

Power plants. . , $ 58,200 $ 33,800 

Tank plants 49,150 41,965 

Melting plants 14,385 22,000 

Slime plants 24,050 16,150 



$145,785 $113,915 

Allowing for land, engineering expenses, shipping facilities, 
etc., total cost may be taken at $2,000 to $3,000 per daily 
ton capacity. The above figures are only intended to serve 
as a basis for computations, and not to furnish exact infor- 
mation, which it is impossible to do anyway as costs are sub- 
ject to great variations, so in many cases I have not thought 
it worth while to try to ascertain exact costs of different 
apparatus. 



CHAPTER VIII. 



PRODUCTS. 



The analyses of refined lead, presented as tables, are col- 
lected from numerous sources, and are not selected in any 
way, but include all the analyses I have. The Consolidated 
Mining and Smelting Company of Canada, Ltd., have kindly 
given me the average analyses of their electrolytic lead and 
lead bullion, which is given as Table 102. 



TABLE 102. 
Bullion. 





Au 


Ag 


Cu 


Fe 


Sb 


1904 averages 


1 . 50 ozs. 
1.00 " 


200 ozs. 
109.1 " 


.50% 
.19% 
.20% 
.20% 


.07% 
• 05% 


.55% 
•44% 
•81% 
•75% 


1905 " 


1906 " 


1907 " so far 















Sn 


As 


Mn 


Zn 


Bi 


1904 averages. . 




.1 Trace 

None 
Trace 


■11% 
.23% 
.15% 
.25% 


Trace 
< < 


Trace 


\one 


1905 " 






1906 " 




< i 


1907 " so far... 





TABLE 103. 
Pig Lead. 



Silver 



Cu 



Fe 



Sb 



Sn 



Bi 



As 



Ni 



Co 



Averages. 



52 ozs. 



0006% 



.0007%[. 0006% 



None None 



None 



None 



None 



284 



PRODUCTS. 



285 



The silver is unusually high in the Trail lead, but with 
other bullion it has probably averaged about i ounce. By 
further washing, the silver may be largely reduced, but they 
find it does not pay to save it.* 

The United States Metals Refining Company, at their plant 
at Grasselli, produce lead of about the following composition :f 







TABLE 104. 








Ag 


Cu 


Sb 


Bi 


Fe 


As 


Pb 


. 00070% 
= .21 ozs. 


.00100% 


.00096% 


.00070 


.00140% 


Trace 


99.99524% 



The quality of the lead varies with the skill and ex- 
perience of the workmen in drawing cathodes and washing 
them. An inexperienced man is apt to wipe off slime from 
the anodes on the cathodes in drawing the latter. The fol- 
lowing data from Trail, 1902, illustrates this: 

TABLE 105. 



Cast. 



Aug. 17. 
19. 
21. 
23. 
25. 
27. 
27. 
29. 
31. 

Sept. 2. 

4. 

7. 

8. 
10. 
12. 
15. 
16. 
18. 
20. 



Oz. Ag 


in Lead. 


0.48 


0.35 


0.26 


0.17 


0.14 


0.26 


0.25 


0.20 


0.32 


0.28 


0.19 


0.25 


0.24 


0.28 


0.29 


0.43 


0.45 


0.39 


0.40 



Cast. 



Sept. 22 
24 
26 

27 
29 



Oct. 



1 
3 

4 
6 
8 
10 
13 
15 
16 
18 
20 
22 
25 
27 



Oz. Ag 
in Lead. 



0.43 
0.35 
0.18 
0.30 
0.32 

0.14 
0.15 
0.13 
0.22 
0.17 
0.17 
0.16 
0.15 
0.10 
0.16 
0.15 
0.11 
0.14 
0.26 



Cast. 



Oct. 28 
30 

Nov. 3 
5 
7 
10 
13 
15 
19 
19 
23 
25 
28 
28 



Dec. 



Average 



Oz. Ag 
in Lead. 



0.24 
0.23 

0.38 
0.34 
0.38 
0.34 
0.35 
0.24 
0.22 
0.23 
0.20 
0.18 
0.21 
0.22 

0.19 
0.12 

0.25 



* Communicated by the Company. f Ditto. 



286 



LEAD REFINING BY ELECTROLYSIS. 



The bullion averaged 310.4 ozs. Ag, and 3.15 ozs. Au. The 
increase in silver about October 27th and November 3d was 
caused by putting on new men at drawing and washing cath- 
odes, who gradually became accustomed to the work, with a 
consequent slow reduction in the silver figures. 

As showing the unequal distribution of silver in the cath- 
odes, the following data by Dr. E. F. Kern are interesting: 



TABLE 106. 



Ag. 



Rough sample from center of steel cathode 97 ozs 

Sample from edge of same rough sheet 1 . 64 

Large warts of lead on steel cathode 2 . 44 

Smoother cathode from same tank . 23 

Smooth and bright cathode . 04 

Smooth heavy cathode . 09 

Smooth deposit on steel cathode 0.07 



TABLE 107. 
Analyses of Refined Lead. Trail, 1902. 



No. 



1 

2 

3 

4 

5 

6 

7 

8 

9 

10 

11 

12 

13 

14 

15 

16 

17 

18 

19 

19 

20 

20 

21 



Cu, 
Per Cent . 



As, Sb, 
Per Cent. Per Cent 



Fe, 
Per Cent 



0006! 

0003 

0009 

0016 

0003 

0020 

0004 

0004 

0005 

0003 

0003 

0005 

0005 

0004 

0003 

0006 

0006 

0006 

0005 

0005 

0004 

0004 

0022 



0.0008 
0.0002 
0.0001 



None 
None 



0.0005 
0.0010 
0.0009 
0.0014 
0.0060 
0.0010 
0.0066 
0.0038 
0.0052 
0.0060 
0.0042 
0.0055 
0.0055 
0.0063 
0.0072 
0.0062 
0.0072 
0.0057 
. 0066 
0.0044 
0.0047 
0.0034 
0.0010 



0.0010 
0.0008 



0.0003 
0.0046 
0.0013 
0.0004 
0.0004 
0.0003 
0.0013 
0.0009 
0.0007 
0.0005 
0.0003 
0.0012 
0.0011 
0.0010 
0.0016 
0.0011 
0.0015 
0.0016 
0.0046 



Zn, I Sn, 
PerCent Per Cent . 



None 



None 



None 



AgOz. 

P.T. 



0.0035 

0.0035 

0.0039 

0.0049 

0.0059 

0.0049 

0.0091 

0.0012 

0.0024 

0.0083 

0.0080 

0.0053 

0.0140 

0.0108 

0.0072 

Trace 

0.0081 



0.24 
0.47 
0.22 
0.22 
0.14 
0.25 
0.28 
0.43 
0.32 
0.22 
0.11 
0.14 
0.24 
0.22 
0.23 
0.34 
0.38 
0.35 
0.22 
0.23 
0.38 



Ni.Co.Cd Bi 
Per Cent. PerCent 



None 
None 



None 



None 



PRODUCTS. 

TABLE 108. 
Analyses of Refined Lead. Trail, 1903 or 1904. 



287 



Silver, 


Copper, 


Lead, 


Iron, 


Antimony, 


Tin, 


Bi.Co.Ni, 


Per Gent. 


Per Cent 


Pei Cent. 


Per Cent. 


Per Cent. 


Per Cent 




00129 


.0015 
.0005 




.0015 
.0015 




.0148 
Trace 


Nil 


.00129 


99.996 


.0006 




.0015 


.0011 


99.976 


.0015 


.0003 


11 




.00030 


.0014 


99.995 


.0015 


.0006 


1 1 




.00192 


.0005 


99.995 


.0017 


.0003 


i i 




.00077 


.0010 


99.997 


.0013 


Trace 


i ( 




. 00084 


.0020 


99.995 


.0015 


.0003 


i c 




.00091 


.0007 


99.996 


.0015 


.0009 


1 1 





TABLE 109. 

Analyses of Refined Lead. Trail, 1904. 
Letter from Mr. W. H. Aldridge. 



Silver, Per Cent. 


Copper, 
PerCent 


Lead, 
Per Cent. 


Iron, 
PerCent 


Tin, 
PerCent 


Anti- 
mony, 
Per Cent. 


Arsenic, 
PerCent 


Bi, 

PerCent 


Zinc, 
PerCent 


.0013 =.38 ozs. 
.0017=. 50 " 
.0019=. 55 " 


.00075 
.001 

.0009 


99 . 9938 
99.9930 
99.9943 


.00075 

.0012 

.0007 


.0001 
.0001 
.0001 


.0028 
.0026 
.0017 


None 


None 


.0005 

.0004 
.0004 



TABLE 110. 
Analyses of Bullion. Trail, 1902. 



No. 


Fe, Cu, 


Sb, 


Sn, 


As, 


Ag, 


Au, 


Pb, 


AgOz. 


AuOa 


PerCent PerCent 


PerCent 


PerCent 


PerCent 


PerCent 


PerCent 


PerCent 


P.T. 


P.T. 


1 


0.00750.1700 0.5400 


0.0118 


0.1460 


1.0962 


0.0085 


98.0200 


319.7 


2.49 


2 


0.0115 0.1500 0.6100,0.0158 0.0960 


1.2014 


0.0086 


97.9068 


350.4 


2.52 


3 


. 0070 . 1600 . 4000 . 0474 . 1330 


1.0738 


0.0123 


98.1665 


313.2 


3.6 


4 


. 0165 . 1400 . 7000 . 0236 . 3120 . 8914 


0.0151 


97.9014 


260.0 


4.42 


5 


0.0120 0.1400 0.8700 0.0432 0.2260 0.6082 


0.0124 


98.0082 


177.4 


3.63 


6 


. 0055 . 1300 . 73000 . 0316 . 1030 . 6600 . 0106 


98.2693 


192.5 


3.10 


7 


. 0380 . 3600 . 4030 Trace 


0.7230 


0.0180 


98.4580 


210.9 


5.25 



288 



LEAD REFINING BY ELECTROLYSIS. 



TABLE 111. 

Slime Analyses. 



No. 



Anodes. 



Cu, 
PerCent 



PerCent 



Sb, As, ! Pb, 

PerCent PerCt. PerCt. 



1 

2 

3 

4 

5 

6 

7 

8 

9 

10 

11 

12 

13 

14 

15 

16 



Lead, Trail, B. C 

Lead, Trail, B. C 

Lead, Monterey, Mexico 

Lead, Mexican 

Lead, Mexican 

Lead, Trail, B. C 

Lead, Trail, B. C 

Rich lead, Parkes process 

Lead, Trail, B. C 

Lead, Trail, B. C 

Lead, Trail, B. C 

Lead from El Doctor Mine, Mexico. . . . 
Copper, Montana converter anodes. . . . 
Copper, Montana reverberatory anodes 

Copper, Boston and Montana 

Copper, Boston and Montana 



8.83 

22.36 
1.90 
9.30 
6.38 
1.40 
6.60 

12.56 
7.10 
7.70 
8.1 
7.82 

41 

18 

57 

53.29 



28.15 
23.05 
32.11 
4.7 
3.90 
31.62 
32.21 
78.45 
29.20 
31.90 
14.6 
2.44 
24 
51.4 
14.80 
12.90 



27.10 
21.16 
29.51 
25.32 
50.16 
35.71 
24.60 
4.12 
30.50 
37.60 
27.6 
75.34 



2.00 
3.30 



12.42 17.05 
5.40 10.62 
9.14 9.05 

44.5810.30 

15.23! 5.30 
4.9l! 9.57 
2 . 20 12 . 60 

3.00 

6.10 10.20 
2.80 12.60 
7.0 16.0 
0.24 12.23 



2.60 5.26 
1.15s Tr. 



No. 



Anodes. 



Bi, 

PerCt. 



S. Fe, j Oz. Au. Se, Te, 
PerCt.PrCt.l PrCr. PrCt. 



Lead, Trail, B. C 

JLead, Trail, B. C 

JLead, Monterey, Mexico 

;Lead, Mexican 

Lead, Mexican 

Lead, Trail, B. C 

Lead, Trail, B. C 

S Rich lead, Parkes process 

9 Lead, Trail, B. C 

10 Lead, Trail, B. C 

11 |Lead, Trail, B. C 

12 Lead from El Doctor Mine, Mexico. . . 

13 Copper, Montana converter anodes. . . 

14 jCopper, Montana reverberatory anodes 

15 jCopper, Boston and Montana 

16 Copper, Boston and Montana 



Nil 
Nil 
Tr. 
.52 

19.74 



Nil 



1.27 
1.12 

.49 

Nil 



SS 



0.81 
1.95 



1.35 



5.70 
1.5511.96 



29.1 



180.33 
81.99 



34.5 



18 
38 



2.0 1.00 
26:1.97 



Remarks.— 1, 2. Trans. Am. Inst. Min. Eng., 1904, p. 182. 3, 4. Trans. Am. 
Inst. Min. Eng., 1904, p. 183. 5. Original. 6, 7. Mines and Minerals, Vol. 25 

<1905), p. 288. 8, 9, 10, 11, 12. Original. 13, 14. Trans. Am. Inst. Min. Eng., 

2904, p. 310. 15, 16. Original. 



PRODUCTS. 



289 



TABLE 112. 

Analyses of Bullion and Refined Lead. Troy, N. Y. 





Ag. 
Per Cent. 


Cu, 
Per Cent. 


Sb, 
iPer Cent. 


Pb, 
Per Cent. 


Bullion 


0.50 
0.0003 


0.31 

. 0007 


0.43 
0.0019 


98.76 


Refined lead 


99.9971 







TABLE 113. 

Analyses of Bullion and Refined Lead. Troy, N. Y. 



Cu 

PerCem 



Bi 

PerCent 



As 
PerCent 



Sb 
PerCent 



AgOz. 
P.T. 



PerCent 



AuOz 
P.T. 



Fe Zn 

PerCent PerCent 



Bullion 

Refined lead. 



0.75 
0.0027 



1.22 



0.936 0.6832 
00370.00250.0000 



358.89 



0.0010 



1.71 

None 0.0022:0. 0018 



TABLE 114. 
Analyses of Bullion, Refined Lead and Slimes. Troy, N. Y. 



Pb I Cu 
PrCt. PerCent 



Bullion 96.730.096 



Refined lead. ....... 

Slimes (dry 

sample) j 9.05 



1.9 



As 
Per Cent 



0.85 
0.00130.00506 



9.14 



Sb 

Per Cent 



1.42 

0.0028 
29.51 



Ag Oz. 
Per T. 



about 275 
9366.9 



Ag 
Per Cent 



680.0027 
0.49 



Fe, Zn, 

Ni Co, 
PrCt. 



Bi 



Tr. 



TABLE 115. 
Analyses of Bullion, Refined Lead and Slimes. Troy, N. Y. 





Pb 
Per Cent. 


Cu 
Per Cent. 


Bi 
Per Cent. 


Ag 
Per Cent. 


Sb 
Per Cent. 


As 
Per Cent. 


Bullion 

Lead 


87.14 


1.40 

0.0010 

9.3 


0.14 

0.0022 

0.52 


0.64 


4.0 

0.0017 
25.32 


7.4 
Trace 
44.58 


Slimes 


10.3 



290 



LEAD REFINING BY ELECTROLYSIS. 



The following analyses by the Osaka Technical Analyzing 
Department, presumably of lead in the Japanese market, prob- 
ably give a good idea of the present quality of commercial 
lead in the world's markets: 



TABLE 116. 
Pig Lead Analyses. 
By the Osaka Technical Analyzing Department. 



Per Cent. 


Selby. 


Trail. 


Smelter. 


English 
Chemical. 


B H.P. 


Enthoven 


Lead 


99.9579 
0.0040 
0.0300 
Trace 
0.0001 
None 
0.0010 
0.0008 
None 

0.0004 
0.0024 
0.0003 
0.0027 


99.9890 
Trace 
None 

Trace 

None 

0.0025 

None 

0.0003 

None 

0.0007 

0.0020 

0.0002 

0.0053 


99.9762 
Trace 
0.0046 
0.0002 

Trace 

< < 

l c 

0.0003 

None 

. 0137 

None 

0.0090 

Trace 

0.0039 


99.9693 
Trace 

0.0007 
0.0003 
Trace 
0.0020 
None 
0.0097 
0.0149 
None 
0.0002 
Trace 
0.0029 


99.9853 
Trace 
None 
Trace 
None 
Trace 
0.0009 
None 

0.0108 

0.0004 

None 

0.0001 

0.0025 


99.9851 


Insolubles 

Bismuth 


Trace 
0.0048 


Cadium 


Trace 


Nickel 




Cobalt 

Silver 


0.0015 


Manganese 

Copper 

Antimony 

Tin 


None 

0.0160 
None 


Arsenic 


« i 


Zinc 


Trace 


Iron 


. 0026 







Note. — Selby and Smelter are American; Trail, Canadian; Enthoven 
and Chemical, English; B. H. P., Australian. 



CHAPTER IX. 
TREATMENT OF LEAD CONTAINING BY-PRODUCTS. 

The refining of copper-lead alloys with high copper is 
of some importance. First, because such alloys can be pro- 
duced as "bottoms" from copper-lead matte, and a method 
of saving both lead and copper is then provided. Second, 
because some lead bullions give a good deal of dross in remelt- 
ing which can not very well be stirred into the lead to make a 
uniform anode, and the natural method of treating such 
drosses, containing as they do from 80% to 90% lead or more, 
is to get them into some kind of an anode and extract the 
lead electrolytically in the usual manner. 

Dr. E. F. Kern tested many methods of treatment in my 
laboratory using an alloy of 60% Pb, 39% Cu, and 1% Ag, 
which methods included removing of the lead by the com- 
bined action of fluosilicic-acid solution and air, the alloy losing 
2% in twenty-one hours. The alloy was also ground up, 
mixed with broken electrolytic lead peroxide, and let stand 
three and one-half days with solution containing lead fluo- 
silicate and fluosilicic acid. At the end of that time all the 
lead had dissolved out, as well as some copper, leaving a 
porous copper material of the same shape as the original 
pieces of alloy. Some Pb0 2 remained, the execution of the 
experiment being faulty in not using the right amount of 

291 



292 LEAD REFINING BY ELECTROLYSIS. 

Pb0 2 to either dissolve the lead alone, or both the lead and 
copper. 

The chemical reactions are: 

Cu + Pb0 2 + 2H 2 SiF 6 = PbSiF 6 + CuSiF 6 + 2H 2 

Pb + Pb0 2 + 2H SiF 6 = 2PbSiF 6 + 2H 2 

This method would not be a promising one, although the 
lead, or lead and copper dissolved, as well as the lead peroxide 
and the fluosilicic acid, could be recovered by electrolyzing 
the solution with metal cathode and carbon anodes. When 
this is done, lead peroxide deposits on the anodes as a hard, 
greenish black, lustrous, well-conducting deposit of a smooth- 
ness superior to most metallic deposits. The reactions are 
the reverse of those just given. The electromotive force when 
depositing lead is about 2.1 volts, and 1.7 volts when deposit- 
ing copper. 

The alloy, or a very similar one, was also treated with a 
solution of ferric fluosilicate which dissolved out the lead, 
and also traces of copper, from using a little too much ferric 
salt. The residue retained the original shape of the alloy, 
but was very soft and porous, consisting of copper and silver. 
In this process the solution was to be electrolyzed for the 
lead and recovery of the ferric fluosilicate. 

Far the best method consists in treating the alloy as anode 
in the usual lead-depositing solution, with a somewhat smaller 
current density. 

In one experiment the following data were noted: Cur- 
rent density about 12.5 amperes per square foot. Distance 
between electrodes 1 to 2 inches. Volts about .15 to .20, rising 
later to .42 volts, when slime had to be removed. Solution 



TREATMENT OF LEAD CONTAINING BY-PRODUCTS. 293 

4% Pb, 15% SiF 6 . The anode was 1 inch thick and the slime 
had to be removed several times before the anode was com- 
pletely decomposed. 

TABLE 117. 

Weight anode 1778 gr. 

Lead deposited 732 gr. 

Slime 340 gr. 

Alloy remaining 675 gr. 

Dr. Kern put 85 gr. of the slime in a small lead box 
with perforations in the sides, and electrolyzed it with a solu- 
tion containing 20% CuS0 4 -5H 2 and 5% H 2 S0 4 , with cop- 
per cathodes. 

As the slime settled down 50 gr. more were added, 
making 135 gr. used altogether. Electromotive force .2 volts; 
copper deposited 100 gr.; weight of residue of silver and 
lead sulphate 37 gr. 

Analysis of these figures indicates the following: 135 gr. 
slime result from about 475 gr. of the alloy. The slime con- 
tains then 4 gr. silver and 33 gr. lead sulphate, per- 
haps a little of this coming from the lead box which lost a 
little in weight. As a result, practically all the copper was 
recovered, but 8.8% of the lead was apparently converted into 
lead sulphate. Perhaps some of this apparent loss was due 
to insufficient washing of the slime. 

In the aggregate the quantity of copper dross converted into 
matte by resmelting, must be quite large, and the lead lost 
in the final conversion of the copper-lead matte is well worth 
saving. Refiners could very well treat their drosses by 
casting into anodes at a red heat and extracting the lead 
electrolytically. 

Experiments on refining hard lead with 18.8% Sb, and 



294 



LEAD REFINING BY ELECTROLYSIS. 



rich lead from the Parkes or Pattinson processes, are described 
in Chapter I. The antimony slime could be refined direct 
with the antimony fluoride solution, as it retains consider- 
able mechanical strength, or it could be cast into anodes. 

The electrolytic lead process ought to be of advantage in 
a small way in some other branches, as for instance refining 
the lead-gold bullion produced in cyanide mills. 

Experiments on galena direct have been fruitless. 

Electrolytic refining of lead bullion high in bismuth, is 
practiced on a small scale, primarily to produce bismuth. 

Table 118 gives the composition of various alloys which 
have been successfully refined, producing pure lead at the 
cathodes. Table 25, on page 68, gives a number of others 
refined by Senn.* 

TABLE 118. 
Analyses of Lead Anodes. 



Pb 



88. % 
82.37% 
65.37% 
65.56% 
82 . 79% 
88.52% 
60 % 
87.14% 



Cu 



Sb 



1.53%? 
2.22%i 

19 

1.94% 18 

• 97% | 9 

• 68% 6 
39.0% 
1.40% 



5 % 

77%! 

51% 5.85% 
24% 5.47% 
12%; 2.73% 
08% 1.82% 



4.10% 



7.40% 



Ag 



9.75% 

14.60% 

1.95% 

1.94% 

• 97% 



1.11% 
.19% 
7.32% 
6.94%, 
3.42% 



.68%i 2.28% 

1-0 % 

.64% 0.14% 



Current Density 
Per Sq. Foot. 



11 



7 
7 
2.0 
2.0 
2.0 
2.0 
4- 7 
.6-17 



amps. 



Slime 

Contains. 



5.3% Pb 



10.3% 



* Zeitschrift. fur Elektrochemie. 1905. Vol. XL, page 229. 



CHAPTER X. 
ANALYTICAL METHODS AND EXPERIMENTAL WORK. 

Slime. — Dissolve 1 gr. in HC1 and KC10 3 , boil out chlo- 
rine, add a little water, neutralize with dry sodium carbonate, 
add excess of Na 2 S solution (prepared from caustic soda by 
saturating with H 2 S, then adding another portion of caustic 
soda of same amount, and allowing to settle before using). 
Heat on plate for an hour or so, filter, add 2 to 3 gr. pure 
caustic soda or potash, and determine Sb electrolytically. 

The following remarks will be useful in making electro- 
lytic antimony determinations. If you are using a smooth 
platinum cathode, deposit on it a layer of antimony from a 
fairly strong solution of tartar emetic to which a little nitric 
acid has been added, and the precipitated Sb 2 3 redissolved by 
adding tartaric acid. Use a current of about 1 to 2 amperes 
per square foot in preparing the cathode, which is then washed 
with water, dried and weighed. 

The antimony deposits from the sulphide solution made 
as above on the prepared cathode in a beautiful, smooth con- 
dition fit for accurate determinations. I usually start the 
electrolysis cold with a current of \ ampere for a cathode 
having 20 square inches of surface, and heat the solution up 
while the current is on to 70° and increase the current to If 
to 2 amperes. After about three hours turn off the heat, and 
after cooling remove cathode, plunge into distilled water with- 

295 



296 LEAD REFINING BY ELECTRLOYSIS. 

out interrupting the current, wash, dry, and weigh. The 
use of alcohol in drying is of no advantage, as the antimony 
does not oxidize very readily anyway. I have found the same 
weight either way. 

Add dilute H 2 S0 4 to filtrate, heat, filter off As 2 S 3 + S, add 
fairly dry paper and precipitate to about 40 cc. concentrated 
HNO4, and digest gently on plate for six to eight hours while 
acid is slowly driven off. This removes small amounts of 
chlorine and all paper. Determine As by Pearce's silver arse- 
nate method, described in numerous books. 

For copper, silver, bismuth, iron, and lead I have taken 
a separate sample, dissolved in nitric and tartaric acids, neu- 
tralized with soda, added Na 2 S digested and filtered. Pos- 
sibly the solution running through is equally suitable for deter- 
mining arsenic and antimony, though several failures, per- 
haps due to other reasons, have always prevented successful 
results so far. 

The insoluble sulphides with the filter paper are dried 
placed in a small beaker, a light applied, when the paper burns 
off and carbonizes. Concentrated H 2 S0 4 is added and gently 
boiled, cover on, till carbon is all gone and solution is clear 
greenish. Possibly sodium or potassium bisulphate would 
work quicker. 

After cooling add water and pass H 2 S. Filter off iron, 
determine it by boiling H 2 S from filtrate, and titrating with 
permanganate. Redissolve sulphides in H 2 S0 4 in same way 
again. Then neutralize with soda and add KCN free from 
sulphide. Pb and Bi remain insoluble as carbonates, while 
silver and copper dissolve. 

The silver and copper cyanide solution, may be acidified, 
AgCN filtered off, and copper determined electrolytically. I 



ANALYTICAL METHODS AND EXPERIMENTAL WORK. 297 

have, however, got good results by electrolyzing the solution 
for silver, using a single dry battery, giving 1.3 volts about 
as a maximum, for source of current. Time required about 
four or five hours, if solution is warm. Then acidify solution 
with nitric acid under the hood, evaporate down, to remove 
the HCN and determine copper electrolytically. Copper and 
silver can also be determined separately, the first by dissolving 
1 gr. of the slime in nitric acid, removing silver as chloride, 
precipitating with ammonia, filtering and titrating with KCN, 
while silver will often be determined in a works by assay. 

The bismuth and lead carbonates obtained as above are 
dissolved in dilute nitric acid, the solution is almost neutra- 
lized with ammonia, heated, and a few drops of HC1 added 
to throw out BjOCl (Ledoux's method *) , which can be dried 
and weighed at 100° in a Gooch crucible. A convenient and 
satisfactory filter for a Gooch crucible consists of a small disc 
of filter-paper, the same size as the bottom of the crucible. 

Add sulphuric acid to the filtrate from BiOCl, and evap- 
orate for lead sulphate, which may be determined in several 
familiar ways. 

The analysis of metallic antimony can be made in the same 
way as the analysis of slime given above, omitting the separa- 
tion and determination of elements known to be absent. 

Assay of dore bullion. — The method in general use in 
the refineries and assay offices of this country is about as 
follows: The determination of silver is carried out by Gay- 
Lussac's method of precipitation with salt, although Vol- 
hard's method, using a standard solution of thiocyanate, gives 
good results unless there is considerable copper present. 

* Low, "Technical Methods of Ore Analysis," page 55. 



298 LEAD REFINING BY ELECTROLYSIS. 

Gold. — 5 gr. are digested in a porcelain crucible about 2 to 
2 J inches high, with one to six nitric acid, until solution ceases. 
The solution is decanted, nitric acid one to one added, and 
boiling continued until gold changes color. It is then washed 
with hot water several times, dried and weighed. I under- 
stand at the San Francisco mint 400 parts of gold are added 
to the assay for gold, and a check made up containing 400 
parts of gold. This is then cupelled with lead and parted with 
acid and weighed, and the surcharge, or silver remaining in 
the gold, determined from the check. 

To determine gold accurately a proof should be run, using 
a made-up alloy containing gold, silver, and copper, in about 
the same proportions as known to exist in the dore, cupelling 
with an equal amount of lead, and parting the button and 
weighing the gold in the same manner. This is done at the 
Philadelphia mint. 

Sampling dore bullion may be done by melting in a graphite 
crucible, stirring well, pouring off; after one-third and two- 
thirds are about poured off, collect a small sample by putting 
small crucibles in stream of metal. Both samples are granu- 
lated separately and assayed separately. If they do not agree 
the bar is melted over again.* 

Analysis of refined lead. — .Five hundred gr. of lead are 
cleaned and hammered or rolled into thin plates, being very 
careful to use a perfectly clean and bright hammer and anvil 
to avoid introducing iron into the sample. The lead is dis- 
solved in a large beaker on the hot plate, in 500 cc. nitric acid 
1.42 and 1,000 cc. water. If the solution gets too hot it will 
foam very much and run over, so that it is necessary to watch 

* Selby Smelting and Lead Company. 



ANALYTICAL METHODS AND EXPERIMENTAL WORK. 299 

it until most of the lead is dissolved. For the same reason roll- 
ing or hammering the lead into very thin strips is not desirable. 

After all the lead is dissolved the solution is generally per- 
fectly clear, although if more than .02-03% of antimony or 
any tin is present, it will show some turbidity.* The solu- 
tion should be diluted to nearly 2 litres to prevent lead nitrate 
crystallizing out on cooling. If not perfectly clear it is fil- 
tered into a 2-litre measuring flask, otherwise it is merely 
transferred thereto. 145 cc. concentrated sulphric acid, pre- 
viously diluted with water, are added, and the flask filled to 
the mark. After settling 1,700 cc. of clear solution are secured 
by pouring through a large filter. 100 gr. of lead as sulphate 
occupy 23 cc, so that we have in solution i|-§~| of the impuri- 
ties in 500 gr. of lead = 451 gr. lead. 

The 1,700 cc. are evaporated to fumes of H 2 S0 4 , taken up 
with 50 cc. water, and the lead sulphate filtered off. The 
lead sulphate is digested with pure sodium sulphide solution, 
filtered and added with the other sodium sulphide solution 
obtained further on. The filtrate from the lead sulphate is 
treated hot with H 2 S for some time and the gas passed 
through until cold. After settling completely it is filtered, 
and iron and zinc determined in the filtrate, while the sulphides 
are treated with Na 2 S. Determine antimony and arsenic as 
described under slime analysis. 

The insoluble sulphides of lead, bismuth, copper, and sil- 
ver may be dissolved in nitric acid, neutralized with sodium 
carbonate, and KCN added. Lead and bismuth carbonates 
are filtered off, the filtrate acidified with H 2 S0 4 under the 
hood, AgCN filtered off and the solution boiled to expel all 



* "Quantitative Chemical Analysis by Electrolysis." Classen-Herrick. 
Boltwood 265. 



300 LEAD REFINING BY ELECTROLYSIS. 

HCN, after which copper is determined in the solution as fol- 
lows: Nearly neutralize the solution with ammonia, keeping 
the bulk small, say 50 cc, add ammonium acetate, and divide 
into two equal parts. Add to one part a fair excess of potas- 
sium ferrocyanide solution, and filter off the red precipitate 
immediately, passing through the paper twice if necessary. 
Add 1 cc. acetic acid to each and the same amount of potas- 
sium ferrocyanide to the unfiltered half, and match the color 
in the filtered half by adding a weak copper sulphate solu- 
tion of known strength from a burette, allowing one minute 
between each addition of copper sulphate, for the color to 
develop.* 

The silver cyanide precipitate is not desired, for silver is 
determined by cupelling a separate sample of the lead. 

To determine bismuth, dissolve the carbonates of lead 
and bismuth in dilute nitric acid and precipitate as BiOCl, 
by Ledoux's method, as described under "Slime." 

To be sure of the results it is necessary to run a check 
analysis on the nitric and sulphuric acids, evaporating the 
same amount of them down nearly to dryness, and treating 
the last of the sulphuric acid in the same way as the lead 
sample. 

The results of refined lead analysis are more apt to depend 
on the chemist than on the lead, and it is desirable that as 
many errors as possible be eliminated to get accurate results. 
One of the causes of error is in the chemicals used, which are 
not absolutely pure of course, and import certain quantities 
of iron, copper, arsenic, and antimony. The amount of nitric 
and sulphuric acid used is as great as the lead sample, so that 

* Crooke's "Select Methods of Chemical Analysis," page 338. 



ANALYTICAL METHODS AND EXPERIMENTAL WORK. 301 

a check should be run on the acids. On one occasion I deter- 
mined copper in lead as .0010%, but on running a check on 
the acid, it was discovered that there was no copper in the 
lead, but it all came from the chemicals. 

The following show the variation of results on the same 
sample of electrolytic lead: 



TABLE 119. 



Fe 


Zn 


Sb 


Cu 


As i Bi 


Ag 


Chemist . 


.00023% 
. 00032%, 


.0004% 


.0007% 


.00043% 
.00045% 
.00045% 

.0013 % 
.0012 % 


. 00005% 

.0065 % 
.0092 % 


None 


.00003% 

.00016%, 
.0003 % 


Betts 


.00040% 
.0022 % 

.0037 % 








.0079% 
.0042% 


.0013% 
.0007% 


New York 

No. 1 

New York 

No. 2 



As the lead was deposited electrolytically and could have 
contained no zinc, the figures by 2 and 3 are certainly wrong. 
My iron and copper determination was made in triplicate and 
all results agreed fairly well, especially for copper. There is 
no agreement at all for arsenic and silver, but I have no con- 
fidence in my own figures for these elements. As it would be 
easy to introduce traces of iron into the sample, unless hammered 
or rolled with care, I think my own figures are nearer right. 
The same remark applies to the presence of iron in the lead 
as to zinc. As the precipitated lead sulphate may take out 
some antimony, the figures for Sb .0007% by two chemists 
are apt to be slightly too low. 

The following analyses were made in the same sample, one 
at Trail, by Dr. Wm. Valentine, and one by Messrs. Ledoux & 
Co., of New York. 



302 LEAD REFINING BY ELECTROLYSIS. 

TABLE 120. 



Cu 


Sb 


Fe 


Sn 


Ag 




.0003%, 
.0020% 


.0060% 
.0010% 


.0003%, 
.0046% 


.0049%, 
.0095% 


.0006% 
.0006% 


Valentine 
Ledoux & Co. 



The agreement in the case of silver is satisfactory. 
The lower figures for iron and copper show less con- 
tamination of the sample mechanically or by chemicals, 
Dr. Valentine's Sn + Sb = 0.105% and Ledoux & Co.'s Sn + Sb 
= .0109%, so that the separation was probably not complete 
in one case. 

Antimony, arsenic, and tin are determined by us in the 
sulphide solutions by electrolysis. Antimony only is removed 
when the solution is elect roly zed. This is an accurate 
method. 

Slag from fusing slime. — This contains antimony, arsenic, 
lead, bismuth, copper, iron, silica, and sulphur. Dissolve in 
HC1, add KC10 3 , boil out chlorine, neutralize with sodium 
carbonate, and determine antimony and arsenic, as in making 
slime analyses. The insoluble sulphides may also be further 
analyzed as in the slime analysis. 

Electrolyte. — To determine acidity, the following method 
was in use at Trail. Add an equal volume of alcohol to the 
sample and titrate with KOH and phenolphthalein, correcting 
for iron and alumina. To determine lead add H 2 S0 4 , filter 
and determine lead by the molybdate method. The following 
method is used in my laboratory: Add alcoholic potassium 
acetate solution. Filter off K 2 SiF 6 , wash with diluted alcohol, 
add paper and precipitate to distilled water in a beaker, heat 
to boiling and titrate with NaOH, using rosolic acid preferably, 



ANALYTICAL METHODS AND EXPERIMENTAL WORK. 303 

but also phenolphthalein as indicator. Load is detennined 
as described above. Also the analysis may be made by 
adding neutral ammonium sulphate, filtering and determining 
lead sulphate. Titrate filtrate with cochineal and standard ^ 
ammonia in the cold. Other determinations on electrolyte 
are seldom made. For free HF, add to hot solution, hot boric 
acid of known strength until a permanent precipitate of silica 
results. 

Reaction : 

4HF + B (OH) 3 = BHF 4 + 3H 2 0. 

Also remove lead with H 2 S filter, let stand till H 2 S has passed 
off or oxidized, and determine HF and H 2 SiF 6 as described 
under the analysis of fluosilicic acid, page 177. 

Copper-silver matte from melting slime. — To determine 
sulphur, dissolve in concentrated nitric acid. All or nearly 
all of the sulphur oxidizes. Dilute and filter. Remove silver 
from filtrate with HC1, add filtrate to insoluble portion, add 
KCIO3 and evaporate to dryness. Add HC1, to dissolve salts, 
then add ammonia until slightly alkaline, and filter. Add HC1 
and BaCl 2 to filtrate. To determine lead, copper, and silver 
dissolve 1 gr. in boiling concentrated sulphuric acid, cool, 
dilute, filter off PbS0 4 and titrate by Alexander's molybdate 
method. Determine silver in filtrate with NH 4 CNS solution, 
filter, add ammonia, filter, and determine copper in filtrate 
with KCN. To determine antimony fuse 1 gr. in porcelain 
crucible with 3 gr. sulphur and 4 gr. sodium carbonate, 
take up in water, filter, add H 2 S0 4 to precipitate sulphides, 
dissolve sulphides in HC1 and KCIO3, boil out chlorine, reduce 
with sodium sulphite, boil out S0 2 and titrate with perman- 
ganate. In determining antimony in the chloride solution by 



304 LEAD REFINING BY ELECTROLYSIS. 

« 
permanganate, the solution should be cool and of consider- 
able volume, and must contain enough HC1 to prevent the 
formation of a brown color on adding permanganate and not 
enough to decompose permanganate fast enough to interfere 
with the end point. In reducing with sodium sulphite, I add 
the sodium sulphite to the solution containing say J to J strong 
HC1, and heat to boiling very slowly to give the S0 2 plenty 
of time to act. Then boil off say J the total volume, cool, 
dilute somewhat, perhaps adding HC1, and titrate. To make 
sure of the result more sodium sulphite and HC1 may be added 
after finishing the titration, solution gradually heated, then 
boiled and titrated again. 

Method of determining silica in slime. — Five gr. of slime 
is treated with moderately strong HN0 3 and boric acid, fil- 
tered, silver precipitated by HC1, evaporated to dryness several 
times with HC1. The residue from the nitric acid was dis- 
solved in HC1 and solution evaporated to dryness several times 
with HC1. Both of these evaporations were taken up with 
HC1 and the insoluble material filtered off. The residue of 
the slime from the treatment with HC1 was treated with aqua 
regia and insoluble material filtered off. All the insoluble 
matter was ignited together, weighed, pure HF added, HF 
and H 2 SiF 6 driven off, and residue weighed again, calling the 
difference silica. 

Antimony fluoride solution. — It is frequently convenient 
to titrate this with permanganate, after diluting the sample 
with water and HC1. If a strong yellow color develops, the 
result is too high, and the proportion of HC1 was not high 
enough, or the sample was too concentrated. The solution 
can be standardized against ferrous iron; 56 parts iron =60 
parts antimony. 



ANALYTICAL METHODS AND EXPERIMENTAL WORK. 305 

Experimental work. — Preparation of fluosilicic acid. Put 
hydrofluoric acid 15-20% strength in a lead pan and add excess 
of finely powdered calcined flint, which dissolves more readily 
than quartz. Heat, but not to boiling, until solution is satu- 
rated with silica, or until pungent smell of HF has stopped 
coming off. To make the lead solution, add the right amount 
of white lead, which ordinarily contains 80% of metallic lead. 
Gelatine or glue is added to the electrolyte in the form of a 
strong, hot solution in w T ater. 




Fig. 66. 



Small electrolytic tanks can be conveniently made of wood, 
and soaked or dipped in hot paraffine for some time. After 
cooling a layer of paraffine can be put on the wood by letting 
it cool inside, then adding paraffine and turning the box in 
various positions. This makes a good tank for refining 
experiments. 

To provide circulation, arrangements for experiments as 
shown in Fig. 66 are convenient. The tanks are rectangular 
and conditions as to current density, voltage, solution, tempera- 
ture, products, etc., may be as exactly determined as on a large 
scale. 



306 LEAD REFINING BY ELECTROLYSIS. 

Conductivity measurements are made with sufficient accu- 
racy for practical purposes, by using a small paraffined box 
about 3 inches square and deep, with two pure lead sheets 
at each end. If the width of the box is known and the 
volume of the solution is measured, the cross-sectional area 
can be calculated. Current is passed through while the 
solution is stirred, and the resistance calculated from the 
ammeter and voltmeter readings. As the polarizing e.m.L 
in depositing lead is under .02 volt, the method is suffici- 
ently accurate. 

In antimony depositing with lead rods as anode, practical 
work may be duplicated with one anode by using a tall par- 
affined box with a full-length anode. The box is about 3 inches 
square inside, and faithfully represents a full-size tank, which 
comprises only a large number of the same units, without the 
intervening walls. 

In experimenting on small quantities of slime it can be 
cooked up with solutions in porcelain evaporating dishes. For 
10 or 20 lbs. large stoneware crocks holding 60-80 litres are 
good. Steam can be turned in through a lead pipe for heating 
and stirring. Antimony fluoride solutions can be handled in 
painted lead tanks or paraffined wooden ones. To roast slime 
with sulphuric acid it is sufficient to spread it on a cast-iron 
plate heated underneath. 

For ferric sulphate electrolysis on a small scale a tank, as 
shown in Fig. 67, taking 25-50 amperes is useful. The 
anodes are cast in lead in a slit cut in a board, and hung with 
two narrow boards from the ceiling. For work lasting only 
a few weeks the lead tank and diaphragm can be soldered 
together with coarse solder, 2 or 3 parts of lead to 1 of 
tin. 



ANALYTICAL METHODS AND EXPERIMENTAL WORK. 307 

For reduction of lead-antimony slags, use a lead pan in which 
the slag is spread out and made cathode with a sheet-lead plate 
for anode just above the cathode. For reduction in the fused 




—■Q 



Fig. 67. 



state lead chloride (from precipitating lead nitrate or acetate 
with NaCl) is melted in a porcelain crucible, and two carbon 
electrodes dipped in. The reduced metal drops from the nega- 
tive electrode to the bottom of the crucible. It is necessary to 



308 LEAD REFINING BY ELECTROLYSIS. 

have the crucible well covered to keep the air out. The slag 
is fed from time to time as reduced. 

An experimental tank for one full-sized lead anode, about 
6 inches wide, 30 inches long, and 42 inches deep, with one 
glass end, was built at Trail, to watch the behavior of the 
anode slime. This was not successful in that particular 
instance because the solution happened to be too dark and 
turbid. 



CHAPTER XI. 

BIBLIOGRAPHY. 

Keith Process. Engineering and Mining Journal, 1878, Vol. XXVI, 

page 26. 
Tommasi Process. Comptes Rendus, 1896, Vol. 122, page 1476; also 

Zeitschrijt jur Elektrochemie, Vol. Ill, 1896-97; 92, 310, 341. 
Glaser, Deposition of Lead. Zeitschrijt jur Elektrochemie, Vol. VII, 

1900, pages 365-369 and 381-386. 
Borchers. Lead Refining with Fused Baths. " Electric Smelting 

and Refining." First English Edition. 
Senn. Zur Kenntniss der Elektrolytischen Bleiraffmation. Zeit- 
schrijt fur Elektrochemie, 1905, Vol. XI, page 229. 
Mennicke. Elektrische Zinngewinnung und Zinnraffmationmit Fluss 

und Kieselflussaure. Zeitschrijt jur Elektrochemie, Vol. XII, 1905, 

pages, 112, 136, 161, 181. 
Whitehead. Electrolytic Refining of Lead, etc. Mines and Minerals, 

Vol. XXV, 1905, page 288. 
Jacobs. Lead and Silver Refining at the Canadian Smelting Works, 

Trail, B. C. British Columbia Mining Record, December, 1904, 

page 410. 
Betts. Electrolytic Lead Refining. "Trans. Am. Institute of Mining 

Engineers," also Electrochemical and Metallurgical Industry, 

August, 1903, page 407. 
Haber. Electrochemical Industry. Vol. I, 1903, page 381. Report 

on Electrochemistry in the United States. Zeitschrijt jur Elek- 
trochemie, 1903, page 390. 

309 



310 LEAD REFINING BY ELECTROLYSIS. 

Ulke. The Electrolytic Refining of Base Lead Bullion. Engineering 

and Mining Journal, 1902, October 11th. 
Betts. Electrolytic Treatment of Electrolytic Slime. Electrochemical 

and Metallurgical Industry. 1905, Vol. Ill, pages 141, 235. 

— Pamphlets for Free Distribution. September, 1901; March, 1904. 
Hofman. Recent Improvements in Lead Smelting. Mineral Indus- 
try, Vol. XI, 1902, page 453; Vol. XIV, 1905, page 421. 

Betts and Kern. The Lead Voltameter. Vol. VI, 1904, page 67. 

Betts, A. G. Electrolytic Process of Refining Lead. (Use of Fluo- 
silicic Acid, etc.) United States, 679824, August 6th; Mexico 
2144, August 19, 1901; Canada 72068, July 2, 1901, assigned 
to Canadian Smelting Works; Great Britain, 1758 of January 
25, 1901; Spain 28516, September 16, 1901, lapsed; Australia 
1205, August 3, 1904. 

— Electrodeposited Lead. United States 713278, November 11, 1902, 

reissue 12117, June 9, 1903. 

— Electrolytic Refining of Lead and Lead Alloys. (Deposition of 

Solid Lead.) United States 713277, November 11, 1902, reissue 
12301, January 3, 1905; Australia 1226, August 5, 1904; Spain 
29567, July 2, 1902, lapsed; Italy 156177, July 23, 1902, lapsed; 
Mexico 2261, July 3, 1902; Canada 77357, September 9, 1902, 
assigned to Canadian Smelting Works; Great Britain 7661 of 1902, 
April 1st; Germany 31374 B, 40 C, April 1, 1902, pending; 
France 320097, August 9, 1902, lapsed; Belgium 162413, April 1, 
1902, lapsed. 

— Apparatus for Refining Lead by Electrolysis. (Compression of 

Deposited Lead.) United States 679357, July 30, 1901; South 
Australia 5354, August 8, 1901; Great Britain, 
lapsed; Germany 134861, July 30, 1902, lapsed. 

— Process of Treating Anode Residues. (With Chlorine.) United 

States 712640, November 4, 1902. 

— Plant for the Electrodeposition of Metals. United States 789353, 

May 9, 1905. 



BIBLIOGRAPHY. 311 

— Process of Treating the Metal Mixture Produced as a By-product 

in Electrolytic Metal Refining Operations. United States 793039, 
June 20, 1905; Mexico filed May 8th, granted July 8, 1905; 
Great Britain 15298 of 1904; Australia 3050, April 27, 1905; 
Canada 94675, March 27, 1905; Germany B39592, pending. 

— Process of Electrodepositing Antimony. United States 792307, 

June 13, 1905; Mexico May 8, 1905; Canada 94674, August 15, 
1905; Australia 3049, April 27, 1905; Great Britain 15294, of 
1904,' July 8th. 

— Electrolytic Apparatus. (Making Ferric Sulphates, etc.) United 

States 850127, April 16, 1907. 

— Apparatus for Refining Lead by Electrolysis. (Copper-lined 

Tank.) United States 803543, November 7, 1905. 

— Electrolytic Process, Using Insoluble Anodes. (Making Ferric Sul- 

phate, etc.) United States 803543, November 7, 1905. 

— Electrolytically Refining Silver. (Methyl-sulphate Parting, etc.) 

United States 795887, August 1, 1905. Canada 94676, March 27, 
1905. 

— Electrolytically Refining Metals. Canada 94676, March 27, 

1905. United States June 18, 1907. 

— Apparatus for Refining Lead by Electrolysis. (Systems of Con- 

tacts.) U. S. No. 827702. 

Kern, E. F., assignor to A. G. Betts. Treating Anode Slime. (Roast- 
ing with Sulphuric Acid.) U. S. No. 863601, November 7, 1905. 

Truswell, R. Anode Mold. United States 823977, June 19, 1906. 

Miller, J. F. Method of Lining Tanks for Electrolytic Work. 
U. S. Patent 857886, June 25, 1907. 

— Casting Metal Sheets. One-half assigned to W. H. Aldridge. 

U. S. Patent 857885, June 25, 1907. 



APPENDICES. 

APPENDIX I. 

Plant of the Consolidated Mining and Smelting Company 
of Canada, Limited, at Trail, British Columbia. 

The pioneer electrolytic lead refinery is that of the above 
company, which is located oa the west bank of the Columbia 
River a few miles north of the international boundary. Trail 
has railroad connection with Rossland, which in turn is reached 
by the Great Northern Railroad, and is also connected with 
the Canadian Pacific system at the north. 

The Trail plant has been operated since 1902, with some 
interruptions for enlargements, and has a present capacity of 
80 tons per day, although the bullion received to be treated 
at present amounts to only about 45 tons per day. It will 
probably be only a short time before sufficient lead will be 
locally produced to keep the plant running at its full capacity. 
The operations and plant have been brought to a high standard 
by the capable management, and many good points have been 
developed that should be noted. 

Power is supplied by the West Kootenay Light and Power 
Company from their plant at the Bonnington Falls, about 25 
or 30 miles north, in the form of three-phase sixty-cycle cur- 
rent, I believe at 22,000 volts. It is transformed to 550 volts 
at a sub-station about a quarter of a mile from the refinery, 

312 



APPENDIX. 313 

and near the smelter. At the refinery power plant there is 
a Canadian General Electric 000 H.P., GO-cycle, 550-volt motor 
directly connected to a 3600-ampcre, 60-110 volt electrolytic 
generator of the same make, which supplies power to the lead- 
depositing tanks. A Westinghouse 165 H.P., 550 volt, three- 
phase motor is directly connected to a 105 K.W., 30-volt, 3500- 
ampere, 580-rpm. electrolytic generator of the same make, which 
at present supplies current to the electrolytic antimony-deposit- 
ing tanks. For power purposes there is a 20 K.W., 125-volt 
direct-current generator that supplies power to the crane. The 
crane uses about 2 H.P. when running. The pumps require 
2-3 H.P., using a three-phase motor, and the centrifugal lead- 
pumps use about 3 H.P. each against a six-foot head of lead. 
Probably the average power in use for power purposes does 
not reach 5 H.P. and the maximum in use is about 12 H.P. 

The tank-room 50 feet wide and 315 feet long is subdivided 
about as follows: Adjoining the south end there is a room about 
18 by 40 feet in which the cathodes are hung and straightened. 
In the main building is first a clear space of about 4 feet from 
the wall; next comes a block of 132 tanks occupying a length 
of about 96 to 97 feet. These tanks are in six double cas- 
cades, 11 tanks long, the highest tanks being 47 inches and 
the lowest 20 inches above the floor, which is level. 

At the low end of these tanks is a small space for solution 
launders, and then comes a row across the building of clean- 
ing-tanks, 7 feet 6 inches long and 6 feet 3 inches wide. Then 
comes a 17-foot clear space with a floor of cast-iron plates. 
This space is used for electrode storage, the electrodes resting 
on small cars, and also for working space. Then there is 
another row of six washing-tanks of the same size, followed by 
a three-foot space for launders and bus-bar connections, and a 



314 LEAD REFINING BY ELECTROLYSIS. 

block of 72 tanks in six-tank cascades which occupies 54 feet 
of the length of the building. Then there is a three-foot space 
for launders and bus-bar connections, followed by 60 tanks 
in five-tank cascades = 44 feet. These latter tanks have not 
yet been used on account of the present scarcity of bullion 
to be treated, though they are going to be put in commission 
soon, while the current flowing will be reduced as long as the 
shortage of bullion lasts. Instead of using less than the full 
number of tanks, all the tanks will be worked with a smaller 
current. Then there is a set of electrode storage racks about 
16 feet long, which occupy the full width of the building, 
excepting the aisles. 

The remainder of the building is occupied by the melting 
floor and contains the lead and bullion kettles and casting 
floor. . In a small side room is the apparatus for making start- 
ing sheets. There is also in one corner a lead-pipe machine. 
The floor is subdivided into first a 25-foot clear space to the 
lead kettles, then the lead kettles take 12 feet. Then there 
is another twenty-five foot clear space to the bullion kettles, 
which are placed toward one side of the centre of the build- 
ing and occupy 12 feet of its length, with a final space at the 
end of about 18 feet. 

The tanks are of four-inch fir with bolts passing through 
the wood and are similar to Fig. 38. Mr. John F. Miller has 
described to me his method of lining tanks.* He uses two 
grades of California asphalt, "hard" and U D" grade. These 
are mixed in the proportions required to give a melting point 
of 45° C. Mr. Miller determines the melting points of the 



* Mr. Miller has applied for United States patents on his tank and 
method of lining it. 



APPENDIX. 



315 



mixtures by molding them into cones about 4 inches high, and 
keeps them in a water-bath of a certain temperature for 
twenty-four hours. If the cone does not show any altera- 
tion in shape in that time, the melting-point is some higher 
temperature, and if it runs at all, the melting-point is some 
lower temperature. 

The seams of the tank are made as shown in the sketch 
(Fig. 68). The tank is placed in various positions, the side 
being treated being of course horizontal. The seam is first 







Fig. 68. 



Fig. 69. 



Fig. 70. 



filled with several layers of asphalt by running along the seam 
with a teakettle holding the hot mixture. After the seams 
are all filled on that side, it is then flooded with an asphalt 
layer about one-quarter inch thick. Two men can line two 
double tanks per day. The tanks in the refinery are giving 
excellent satisfaction. Though in use two years they have 
not yet required any repairs. There is little or no absorption 
of the solution by the wood, as is the case with merely painted 
w r oocl tanks. 

There are two styles of bus-bars in the refinery, as shown 
in Fig. 69. They and the cathode cross-bars ' are kept scrupu- 
lously polished so the contact losses are only from .01 to .05 



316 LEAD REFINING BY ELECTROLYSIS. 

volt for each contact, averaging about .02 volts or less. There 
are three sets of contacts to the tank, bus-bar to anode, 
cathode to cathode cross-bar, and cathode cross-bar to bus- 
bar. Small iron clips are driven on the cathode and cathode 
bar, after the tank has been loaded, to ensure good contact 
throughout the tank. The clips are as shown in Fig. 70. 

The anodes are cast in closed upright molds, ten at a time, 
similar to Fig. 71, which shows some new molds that have 
been ordered differing from those at present in use only in 
that they are to be made of steel instead of cast iron, the taper 
allowed for withdrawing the lead is to be less, and the size 
of the anode head is reduced. Mr. Miller has applied for a 
United States patent on this mold. 

The main body of the mold only is to be made of steel, 
and the wedge is to be of cast iron. The present molds ope- 
rate very well, and when the anodes are lifted by power three 
or five at a time, instead of by a chain-block as at present, the 
cost for labor per ton cast is not expected to exceed 20 cents, 
though at present it is higher, namely about 27 or 28 cents 
per ton, with wages of 35 cents per hour. The molds are 
placed upright in a wood box arranged with a set of water 
sprays to cool each mold. The lead is pumped from the kettle 
into the mold with a centrifugal pump, which was originally 
a water-pump. This pump remains at the bottom of the 
kettle continuously and has already been in use for a long 
time with no repairs or cleaning. It. is driven by a 3-H.P. 
motor through a belt and gearing. See Plate 10. 

The anodes weigh 370 to 380 pounds each. At present 
the percentage of scrap returned to be remelted is about 20 
but this will be considerably reduced with the new molds, 
which will make smaller lugs. 



APPKNDIX. 



317 







fe*lH 



318 



LEAD REFINING BY ELECTROLYSIS. 



The anodes are lifted by man power with a chain-block 
and stacked in a vertical position, with the same spacing as 
is used in the depositing-tanks, in cars holding ten anodes 
each. There are 40 of these cars at the plant. Before lifting 



Fig. 72. 

into the tanks two cars are run together, and a spacing-board 
protected from wear by sheet iron is placed over the top to 
give the exact spacing. (Figs. 72 and 73.) 

The anodes remain in the tanks eight or nine days when 
the full current is passing, giving two crops of cathodes four 




Fig. 73. 



to five days old. The production of only one crop of cathodes 
per set of anodes has been tried and it is contemplated to 
go back to it, from which one would conclude that the cost 
of refining was about the same either way. 



APPENDIX. 



319 



The anode scrap with most of the slime still adherent is 
lifted by the crane, a portable tray (Fig. 74) is hung under- 
neath the load by the "crane chasers," when the crane car- 
ries the whole to one of the large washing tanks, where it is 
deposited, while the scrap is handled therefrom individually 
on a chain-block by the men who clean scrap. This requires 
three men on one shift. The cleaned scrap is thrown on a 
small flat car and wheeled to the bullion-kettle and dumped 
in. 

The cathodes are made on the sloping table, one man 




-^O 



Fig. 74. 



making 400 sheets in eight hours, while the night watchman 
makes 200 to 250 sheets during the night while not engaged 
in his other duties. The sheets are taken on small flat cars 
to the hanging room, where they are placed on a table, flattened 
out, wrapped around the cathode cross-bars two or three 
times, the men using a stick to bring the lead close to the 
copper. They are then hung, 21 to the load, on small cars 
provided with convenient supports. To place them in the 
tanks they are wheeled through the aisle to a position oppo- 
site the tank, and a man stands over the tank, reaches over 



320 LEAD REFINING BY ELECTROLYSIS. 

to the car and lifts the sheets one at a time into the tank. 
The entire operation of charging a tank with cathodes, fixing 
the spacing and contacts, and wheeling the car to the tank 
and away again requires about fifteen minutes for one man. 
For lifting cathodes to and from the tank, two styles of lift- 
ing racks are used. There are several of each at the refinery. 
The cathode lifting rack, which is the most complicated, may 
be seen in Plate 11. 

The finished cathodes are lifted by the crane, the same 
pan being placed underneath as before, when the load goes 
to a washing tank where it is deposited and any slime wiped 
off and the plates splashed to get off the strong solution. 
They are then placed by the crane on a portable rack near the 
melting pot, and are pushed over into the pot by hand, which 
only takes a minute or two, the cathode cross-bars of course 
being previously pulled out. 

The cathodes are slowly melted down during the day, and 
on the evening shift the pots are skimmed, and the centrifugal 
pumps lowered into the lead and the lead molded. The lead 
launder is about 22 feet long and there are some 160 molds 
in the circle. A crew of four men can mold about 20 tons 
per hour, but they do not work as fast as this, as the two 
men who wheel the lead to the box-cars and pile it in them 
could not keep up, so that about 15 tons per hour is the usual 
speed. The six men do all the work, including firing, skim- 
ming, and loading. 

The use of a centrifugal pump for raising lead gives the 
"best satisfaction, and is to be preferred to any of the other 
methods. This idea originated with Mr. Miller, of the Trail 
Company. A two-inch pump is about the right size, and costs 
about SI 3. 00 at Seneca Falls, N. Y. It is necessary to find 



APPENDIX. 321 

the right speed at which to run the pump before it can be 
worked satisfactorily. 

The slime is collected from the electrolytic tanks, after 
siphoning off the clear solution, by a man who gets into the 
tank with a pail and shovel, and raises the slime by hand 
into a copper tank about 15 by 30 inches which runs on a 
small car between the tanks. A piece of oilcloth is hung 
over the top of the tank and of the copper tank to keep from 
losing slime and getting the bus-bars dirty. No pains are 
taken to get the tank entirely clean, but on the contrary con- 
siderable slime is usually left in the tank. Quite a little slime 
is also collected from the large washing tanks, a hand-pump 
being used to raise the slime into the copper-tank cars. There 
are six of these cars at the plant, and two men are employed 
cleaning tanks and taking the slime to the slime washing 
tanks. The slime cars are hoisted on an elevator and run 
on rails over the slime washing tanks. A plug at the bot- 
tom of the tank cars is raised with a copper wire, when the 
slime drops through a screen into the washing tanks. Fre- 
quently a hose is turned into the tank car to wash out the 
heavy slime. There are four of these washing tanks, which 
are of wood, side by side, each about 42 inches wide, 8 feet 
long, and 5 feet deep. The decantation method of washing 
is in use, and the results are reported on a slip like that shown 
below. The slime is stirred once by a paddle, and steam blown 
in. After settling, the clear solution is siphoned off into one 
of three launders according to destination, the strong solu- 
tion being returned to the electrolytic tanks, that of medium 
strength going to the evaporators, while the weakest is used 
for washwater. The solution with which the slime is satu- 
rated is finally reduced to about 2° Beaume. The slime is 



322 



LEAD REFINING BY ELECTROLYSIS. 



finally run out through a large hose fastened into one end 
of the tank, and ordinarily held up against the end, into sev- 
eral wood suction-filters. 

TRAIL REFINERY. 

Slimes Washing. 
Date, July 11, 1907. 





Tank No. 4 




Tank No. 5 




Wash No 


Beaume. i Destination. 


Wash No. Beaume. 


Destination. 


Slimes 
Water. : . 

1 

2 

3 

4 

5 

6 


30 

20 

10 

6 

2 




Pump 

Evaporator 

Washing 


Slimes 
Water. . . 

1 

2 

3 

4 

5 

6 


40 

30 

20 

10 

6 

2 




Pump 

i c 

Evaporator 

i < 

Washing 










Tank No. 1 


Tank No. 


Wash No . 


Beaume. 


Destination. 


Wash No. 


Beaume. 


Destination. 


Slimes 
Water. . . 

1 

2 

3 

.4 

5 


40 
22 
20 
10 
5 


Pump 

1 1 
Evaporator 


Slimes 
Water. . . 

1 

2 

3 

4 

5 

6 


30 

20 

10 

5 

2 




Pump 

Evaporator 
Washing 


6 





















Remarks : 



Signature. 



APPENDIX. 323 

The evaporation of the washwater is conducted in a wood 
tank about 8 feet square, in which is dropped a lead pipe 
through which steam is passed. The acid lost during the 
evaporation is three pounds SiF 6 per ton lead refined. 

Slime treatment. — Mr. Alexander McNab's method of treat- 
ing slime is now used. The slime from the washing tanks 
is first sucked dry as possible on the suction-filters. The 
slime is then neutralized by stirring into it a little caustic 
soda, and transferred in about 600-lb. lots to one of six 
or eight large iron tanks about 3 feet wide, 8 feet long, and 
4 to 5 feet deep. The tank is then nearly filled with the so- 
dium sulphide solution as it runs from the antimony deposit- 
ing tanks, and 25 lbs. of sulphur is added. It is stirred 
with a wood paddle once, and steam turned in, which is 
thereafter sufficient for the stirring. After about two hours' 
boiling the solution is settled and siphoned off, and the tank 
is again filled with sodium sulphide, sulphur being omitted 
this time. After boiling and settling again the clear solu- 
tion is siphoned off and added to the same storage tank as 
the first lot. The slime is drained and treated further as will 
be described below. 

The sodium sulphide extracts about 80% of the anti- 
mony and some arsenic, and converts a good part of the re- 
maining metals into sulphides. Contrary to what would be 
expected, most of the arsenic remains in the slime until the 
final melting to dore bullion. The sulphide solution con- 
tains about 3.5% of antimony after the slime treatment, and 
varying quantities of sulphide, polysulphide, and thiosulphate. 
The solution is collected in suitable storage tanks, and run 
through a series of antimony depositing tanks of iron, with 
sheet steel cathodes and lead anodes. The anodes are the 



324 LEAD REFINING BY ELECTROLYSIS. 

same kind of sheets as are used in the lead depositing tanks 
for cathodes. There are ten of these tanks in two cascades 
of five each. They have each about 240 square feet of anode 
and 240 square feet of cathode surface, and can take a cur- 
rent of 3000 to 3500 amperes with a potential when working 
without polarization of about 1.5 volts each. A tempera- 
ture of about 60° C. is used as giving the highest efficiency. 
The iron tank is itself connected as cathode. The lead anodes 
remain in working condition about ten days and are then 
renewed. The current efficiency is about 45%. No dia- 
phragms are used. The solution running out of the last 
tanks contains about 1% antimony, and is returned to be 
used for extracting antimony from fresh slime. The opera- 
tion of the tanks is entrusted to three men, one on each shift. 
When the antimony deposit of the cathodes gets about one- 
eighth inch thick, they are taken out one by one and the anti- 
mony knocked off by hammering. Each tank has about 
20 cathodes about 2 feet wide and 3 feet deep. The sodium 
sulphide is changed by electrolysis to thiosulphate, which 
means a heavy loss of sodium sulphide. Attempts will be 
made to crystallize out the thiosulphate of sodium and re- 
convert it to sulphide by reduction with carbon at a red heat. 
This part of the plant is running at a loss at the present time, 
partly on account of a drop in the price of antimony, but 
mainly because the percentage of antimony in the bullion has 
recently dropped to less than one-tenth of one per cent, and 
the percentage extracted by the sulphide solution is much 
less when there is only a little antimony present. The anti- 
mony deposited contains a little arsenic, which can be re- 
moved if too much is present by melting under an alkaline 
slag. 



APPENDIX. 325 

One of the principal items of cost of the process is the 
heavy cost for sodium sulphide, which is quite expensive when 
delivered at Trail. Mr. McNab mentioned that the loss of 
Xa_>S would be about 30 lbs. or less per ton of Trail bul- 
lion, without recovery or regeneration of the solution. 

To use this process continuously it would be very necessary 
to have extremely good ventilation where the sulphide solu- 
tion is stored and handled, for the gases given off are inju- 
rious. 

The treated slime has run as high as 27% arsenic and 4 
to 10% antimony, while the raw slime contains perhaps 10% 
arsenic. 

The deposited antimony contains some gold and silver. 

The slime is next dried in a large iron pan placed over 
the roasting-furnace flue, with a hole in the bottom, so that 
it can be dropped directly into the furnace. The roasting- 
furnace is of the muffle type, and is hand .raked. Its length 
for the hearth appears to be about 20 feet and width about 
7 feet. The slime is calcined at a very low heat for the pro- 
duction of the oxides, arsenates, antimonates, and sulphides 
of lead, silver, and copper, and is finally raked into a large 
steel bucket suspended by a chain block from an overhead 
runway. The roasted slime is leached with sulphuric acid 
and water, taking out most of the copper and one-third to 
one-tenth of the silver, which is precipitated by metallic 
copper. The copper sulphate is crystallized out and 
sold. 

The residue is melted in a magnesia-line reverberatory 
furnace, using Crow's Nest Pass coal, with silica as a flux to 
slag off the lead. This is a tedious operation, as the lead sul- 
phate and silica do not react readily. They are going to try 



326 LEAD REFINING BY ELECTROLYSIS. 

my suggestion to put some old slag in each charge to help 
the melting. The parting is done by the sulphuric acid 
method, and the copper sulphate is crystallized for the mar- 
ket. They will probably try another suggestion to heat the 
roasted slime with sulphuric acid direct to make nearly all 
the silver soluble, which should save melting the silver 
twice. 

There is a fluosilicic-acid plant which distils a mixture of 
fluorspar, silica, and sulphuric acicl in iron pans about 8 feet 
diameter. The acid fumes are condensed in wood towers 
about 1 foot square and perhaps 20 feet high, through which 
a spray of water is dropped. The fumes pass up and down 
through a series of some six towers. The plant was not in 
operation on account of shortage of sulphuric acid, at the 
time of my visit. Excellent results are claimed for this 
plant. 

Labor required. — The tank-room labor is subdivided as fol- 
lows: In addition to the general foreman, there are three shifts 
of three men each who inspect the tanks for short circuits, clean 
bars, empty and fill tanks with solution, put on and take off 
clips, and clean cathodes. Two men are employed putting 
cathodes into the depositing tanks. Three men clean the 
anode scrap; two men clean tanks; one man runs the crane; 
one man attends to the changing of the electric connections 
and the siphons; three men hang sheets; two boys clean 
cathode cross-bars; one man makes sheets; one man for 
night watchman who also makes some sheets; one man on 
the clay shift and one man on the evening shift are employed 
in washing slime free from lead-depositing electrolyte, and 
two men follow the crane, making one foreman and twenty- 
nine men in all. The wages are 3-i.o cents per hour and the 



APPENDIX, 327 

men get out 45 to 50 tons at the present time in about G.5 
hours. I am informed both by the superintendent and the 
foreman that the same crew could handle the full 80 tons in 
eight hours, with two or three additional men. The reason 
for this is that there is not sufficient work for the men at pres- 
ent, and the men are probably allowed to waste a good deal 
of time, as they are paid by the hour, and labor is so scarce 
that they would leave if they only got five hours work a 
day. 

With a production of 50 tons per day the labor cost is evi- 
dently about SI .40 per ton at present, and with 80 tons pro- 
duced* per day, it would be about SI. 17 per ton. With labor 
at 25 cents per hour it would evidently be for 80 tons per 
day about SO. 85 per ton. Eventually the plant will probably 
have an anode-wiping rig that will handle a tank-load of scrap 
at a time, and if the plant had more slime washing tanks, or 
if they were larger, one man could easily do the work that 
now takes two men, which would reduce the cost per ton at 
Trail on the 80-ton scale by about 10 cents. 

The labor loading and unloading lead and bullion and 
firing and melting takes, in addition to the general foreman 
six men casting anodes (divided into two shifts) and six men 
on one shift casting and loading lead, and five men unload- 
ing bullion and shifting anodes, while one man fires the pots 
and dumps cathodes in the daytime. This includes the sam- 
pling of the bullion and the remelting of the anode scrap and 
skimming the pots. This force is fully employed to handle 
50 tons on an eight-hour shift. The wages are the same as 
for the tank-room force, so the labor cost with the present 
arrangement of plant is evidently about SI. 00 per ton refined, 
including loading and unloading, firing, sampling, and wheel- 



328 LEAD REFINING BY ELECTROLYSIS. 

ing lead and bullion about the refinery. Certain reductions 
in this item are planned. It should be remarked that the 
refinen' has no electric or other power traction system for 
moving lead around, and there is a chance to make a saving 
there. Forty poimds of coal are consumed per ton of lead 
melted. The refinery will probably ultimately receive bul- 
lion in the form of anodes instead of pigs, which will save 
quite a little expense. The repair item is very small with the 
steel pots in use, which last for a very long time. There are 
employed at the refinery two carpenters and one machinist, 
for repairs and improvements. There are two SO-H.P. 
boilers which are fired by the same man who rims the elec- 
tric generators, three in all for the three shifts. 

In the slime plant there are employed one foreman: two 
men boiling slime with sodium-sulphide solution: three men 
on the antimony depositing tanks, who also take care of the 
lead fluosilicate-solution evaporators: two men drying and 
handling slime: three furnacemen on the roasting furnaces: 
and one man in the copper-sulphate crystallizing plant, twelve 
men altogether. I did not make any inquiry about the part- 
ing process and operation, as that is such a well-known process 
anyway. 

The superintendent's assistant keeps the records of opera- 
tion, shipments, etc. 

At the time of my visit the electrolyte in the lead deposit- 
ing tanks was rather weaker than usual owing to a scarcity 
of sulphuric acid for making fluosilicic acid, and contained 
about 5 gr. lead and 10 gr. SiF 6 per 100 cc. I was informed 
that the greatest economy at Trail, after taking into con- 
sideration everything, as power, acid loss, etc.. was reached 
with a solution containing about 12 gr. SiF € per 100 cc 



APPENDIX. 



329 



It would be expected that at Trail, with expensive acid and 
not very expensive power, the greatest economy would be 
achieved by economizing in acid at the expense of some power. 
The acid loss for the proceeding two months had been 7 and 
(i lbs. of SiF 6 per ton lead respectively. Mr. Miller in- 
formed me that he thought it averaged about 8 lbs. when 
the plant was running full. The circulation is maintained 
at about 5 to 7 gallons of solution per minute for each tank. 
The current efficiency averages about 88%. The e.m.f. per 
tank is about 0.4 volts. 

The daily report is made out on the form shown: 



TRAIL SMELTER 
Lead Refinery Report. 



May 31, 1907. 



TANK ROOM. 



Pig Lead Produced lbs. Last 10 Days 465.36 tons 

Pig Lead Produced this Month to Date 1527. 51 tons 

Pipe 34 tons 



[Acid 9.2 10.1 Pet. 

Electrolyte Lead 4.4 5.0 Pet. 

1 Sp. Gr. 1 . 13 1 . 16 

[Temp. 34° C. 



| Average Amperes . . 3066 . 6 
! Average Volts. ..... 72 . 2 

JH.P 296.5 

[Time Running 24 hours 



First Crop. Second Crop. 

Tank Efficiency 95.7 Last 10 Days 98.3 Last Month 86.3 

Lead per K.W. Hour 20.2 " 10 " 20.7 " 



No. of 

Tanks 

Charged. 



Weight Anodes. 



Weight Cathodes. 



Weight Scrap. 



Pet. Scrap. 



Starting Sheets, Made No. 378 Day Shift 



210 



Night Shift 



330 LEAD REFINING BY ELECTROLYSIS. 



MELTING-ROOM. 

Lot No. 1116 967 Pigs 87,915 lbs. 

Refined Lead Shipped Lot No. 1117 435 Pigs 40,001 lbs. 

1402 Total 127,916 lbs. 

|| Pigs lbs. 

Refined Lead on Hand Lead Pipe lbs. 

j Cathodes ibs. 

IITotal lbs. 

Bullion Received, Lot No Bars lbs. 

Bullion on Hand, Trail Bars H. M. Bars 

No. Anodes Made Night Shift, 150 Day Shift, 190 Total 

Remarks 



N. B.— Lead produced does not include pipe or dross. 

Plates 8 to 13 show interior and exterior views of the 
refinery. 




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APPENDIX II. 

Lead-repining Plant of the United States Metals 
Refining Company at Grasselli, Lake County, 
Indiana. 

The principal buildings are an office building, tank and 
melting building, 72 feet by 360 feet for the depositing tanks 
and melting furnaces, a power plant at one end of the large 
building, and a hydrofluoric and fluosilicic acid building. 
The slime-washing machinery and evaporators are located 
in separate small buildings. 

The power plant has two boilers, only one of which is re- 
quired at a time. The fuel is local bituminous coal of fair 
quality. One cross-compound Nordberg engine drives a 
Crocker- Wheeler electrolytic generator, having its maximum 
efficiency at 60 volts and 4,500 amperes and capable of carry- 
ing considerable overload. The same engine drives by belting 
a small generator for power and lighting purposes. A smaller 
double-expansion Nordberg engine drives a Westinghouse 
110-K.W., 110-volt, 1000-ampere dynamo which can be used 
for power and lighting purposes, and has very considerable 
reserve capacity. The power plant and the tank and melt- 
ing plant are in handsome substantial brick buildings. 

The tanks occupy the rear end of the large building near 
the power plant. The tank arrangement follows the Walker 

343 



344 LEAD REFINING BY ELECTROLYSIS. 

system as used in copper refineries, but the number of tanks 
per block is four only instead of the large number used for 
copper. 

Each tank takes 26 anodes (for size of tanks, etc.. see 
page 214) and 27 cathodes of sheet lead weighing about 18 
lbs. each. The cathodes are cast of the form shown in 
Fig. 57, and the sheets are hung over the cathode-bars, which 
are of copper about fxlj inches in cross-section, on a special 
table provided for the purpose. A hole is punched through 
the sheet and the overlapping strips, and the burr produced 
hammered out, giving a satisfactory hold, and a double thick- 
ness of lead at the solution line, which latter is a help in that 
there is little or no chance of the solution cutting through 
the cathode at the surface, during the time the sheet remains 
in the tank. 

The tanks are lined with an asphalt um mixture. Great 
cave is required in getting a proper mixture; one that will 
not soften at the temperature of the electrolyte and will not 
crack in cold weather if the tanks are empty. 

Electric motor-driven centrifugal pumps raise the solution 
from the pump-tanks beneath the level of the depositing-tanks 
to the feed-tanks at a level higher than the depositing-tanks, 
leaving the rest of the flow through the tanks to be accom- 
plished by gravity. The solution circulates through two tanks 
only before it again flows down to the storage and pump 
tanks. 

The tanks are supported on concrete piers, w r hich are well 
asphalted. The ground under the tanks slopes to sumps and 
is also well asphalted. 

Two electric travelling cranes, 72-foot span, capacity 10 
tons, command the entire tank and melting space. One crane 



APPENDIX 345 

could probably do all the work quite well if the other should 
be cnit of order. 

Tapering tank bus-bars are used to save in copper. All 
electrodes are supported on small triangular copper rods fas- 
tened to the bus-bars on the outside of each block of tanks, 
while for the intermediate supports the triangular pieces 
suffice. 

The bullion comes to the refinery already cast in anodes, 
from the United States Smelter near Salt Lake City. They 
are unloaded from the box-cars and sampled by punching, 
with the help of a chain block and a temporary track run 
into each car, at a labor cost of probably about 6 cents per 
ton. The anodes are 2 feet wide and 3 feet deep and weigh 
about 450 lbs. each. 

There are two melting kettles at one end of the main 
building nearest the railroad track, one of which is used for 
melting refined lead and the other for making fresh anodes 
from the anode scrap. The pots are at quite an elevation 
above the floor, so that the lead may be siphoned out, though 
it is the intention to use a centrifugal pump as at Trail. The 
lead is molded in the usual manner and goes into the market 
marked "electrolytic." The bullion is molded into ten flat 
open molds, and removed with an air hoist running on an 
overhead track. 

The washing of the cathodes is now done with a spray, 
though the method in use at Trail will probably be adopted, 
as it is perhaps a little quicker. The anode scrap with attached 
slime is hung by the tank-load in a tank of about the same 
size as the electrolytic tanks, and a gang of five men with scrub- 
bing brushes attached to poles about six feet long, reach in 
between the anodes and wipe off the slime into the solution 



346 LEAD REFINING BY ELECTROLYSIS. 

or washwater in the tank. The crane then picks up the 
load, when it is washed with a spray of water and is then 
carried to the pot, and lowered part way in. When the crane 
travels off the side of the pot draws the cathodes off the lift- 
ing rack, and the cathodes fall in. The crane has two lifting 
ropes, one at each end of the lifting rack, otherwise this 
method would not be practicable. 

The slime removed from the anode scrap, and that col- 
lected from the bottoms of the electrolytic tanks is piped 
to a separate building. Part is pumped into a large iron filter- 
press until the press is filled up, when an air blast is turned 
in to get as much of the strong solution out as possible. The 
slime is then washed with cold water, until the solution 
running out is reduced to 2° Beaume, when the air blast is 
again turned in to dry the slime. The rest of the slime 
is washed in two centrifugal machines with copper baskets. 
The slime is next dumped into iron drying pans heated by 
a fire (steam drying is too slow), and when the moisture is 
reduced from about 40% as it comes from the filtering machines 
to 10 or 20%, it is barrelled and shipped to the company's 
refining-plant at Chrome, N. J., for further treatment. 

The strong solutions and washwaters from the filtering 
plant are probably returned to the electrolytic tanks, while the 
weaker are evaporated. The evaporation is partly carried 
out in wood tanks as at Trail, and also in a large circular 
tank of hard lead, the latter being decidedly the best. 

The lead-depositing electrolyte at the time of my visit 
contained about 6 grams of lead and 9 gr. of SiF 6 per 
100 cc. The temperature was about 32° C. and the volts 
per tank about .45. The solution will undoubtedly be 
strengthened up later. 



APPENDIX. 347 

The acid-making plant is very complete. The fluorspar, 
slightly in excess, is mixed with not too strong sulphuric acid 
and distilled, and the hydrofluoric acid produced is saturated 
with silica in tanks with mechanical agitators. The results 
are excellent, and the building is usually free from acid 
fumes. 

Plates 14, 15, and 16 are views of the works. 




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APPENDIX III. 

Treatment of Lead Refinery Slime with Solution of 
Ferric Fluosilicate and Hydrofluoric Acid. 

The treatment of lead refinery slime is on a fairly satis- 
factory basis, by methods discussed in Chapter II, but iri the 
endeavor to carry the electroyltic treatment to greater per- 
fection I made experiments in my laboratory, which I shall 
describe below. 

The experimental operation was on a scale corresponding 
to the treatment of the slime from one ton of lead bullion 
per day. The experimental plant was run continually twenty- 
four hours each day in charge of two shifts, while daily analyses 
of important products were made, to follow the operation as 
closely as possible. 

There were so many difficulties to contend with, prin- 
cipally with the apparatus, that after running about a week, 
I was obliged to shut down and make changes. After 
starting up again the plant was operated continuously for 
two weeks, until the supply of slime on hand was practically 
used up. 

During the middle of the second run the supply of lead 
electrodes was used up, and the operating force was too busy 
to make more, so that it was a case of shutting down and 
beginning over again, or changing. The deposition of the sepa- 
rate metals was not being done well, and it seemed impossible 

355 



356 LEAD REFINING BY ELECTROLYSIS. 

to do it, at least with the arrangement of plant. Partly under 
the force of necessity, I introduced a change in the process at 
this time, that turned out in a very gratifying way. 

The scale of operation and the desire to operate the plant 
continuously afforded a better test of what could be done on a 
commercial scale, than smaller laboratory tests could have 
possibly given. 

The process used was referred to in Chapter II, page 92, 
and is similar in some respects to processes mentioned on 
pages 119-123 and 134-137. In a general way the process 
consists in attacking the fresh unoxidized wet slime with a 
solution of ferric fluosilicate and hydrofluoric acid which re- 
moves over 99% of the arsenic and copper, 90% or more of 
the antimony, and nearly 90% of the lead. Originally it was 
intended to remove from the resulting solution, first the copper 
as cathode metal, while antimony anodes dissolved, thus 
substituting antimony for copper. Next the antimony was 
to be deposited using lead anodes which dissolve, so that lead 
takes the place of antimony in the solution. After this the 
arsenic was to be plated out as a lead-arsenic alloy, while lead 
anodes were also used in this case, thus substituting lead for 
arsenic. The solution now containing only lead and ferrous 
fluosihcates was to be electrolyzed for metallic lead and ferric 
fluosilicate, the latter to be used over again in the same way 
as before. Of course the different electrolytic steps were to 
be performed in separate sets of tanks through which the 
solution flowed in series. 

Later, the electrolytic deposition of copper, antimony, and 
lead-arsenic was given up and the metals cemented out in 
layers of different composition by causing the solution to flow 
through the lead product obtained in the ferric-iron producing 



APPENDIX. 357 

tank. This gave a separation, although the antimony and 

arsenic were recovered together. 

The use of hydrofluoric acid in the solution is important, 
because without it only a little antimony could be dissolved. 
If hydrofluoric acid only was used lead could not be extracted. 
By the addition of hydrofluoric acid to the solution within 
certain limits, the extraction of the antimony may be secured, 
without spoiling the lead extraction. 

The ferric fluosilicate-hydrofluoric acid seems to me to 
be probably the best of all wet slime processes, because it 
offers these advantages: A minimum of electrolysis to produce 
the desired products; recovery of any lead refining solution 
or any fluorine left in the slime by incomplete washing, or 
decomposition of the lead depositing electrolyte; treatment 
of wet, raw, and imperfectly washed slime; simplicity; no 
chance to lose valuable metals; elimination of arsenic from 
the slime-treating solution; and recovery of the arsenic. Very 
few of the slime processes have any of these advantages. 

Experiments were made to find a suitable electrolytic 
diaphragm capable of withstanding solutions containing hydro- 
fluoric acid. With asbestos and earthenware impossible, it 
is not an easy matter to produce a diaphragm. Quite satis- 
factory tests were obtained with carbon buttons, prepared by 
mixing powdered charcoal with asphaltum varnish and stamp- 
ing into buttons 1J inches diameter and i 3 « inches thick, which 
were dried and baked gently. Using soft charcoal and baking 
below a red heat the product was electrically non-conductive, 
and after removing air under an air-pump, or boiling in a 
solution of sodium nitrate, gave a fair electrolytic conduc- 
tivity. 

Preliminary tests were made on two lots of lead slime. 



358 



LEAD REFINING BY ELECTROLYSIS 



Lot 1 had been partially dried in the usual course of treatment 
and was pretty well oxidized. Lot 2 had been merely filter 
pressed, and was practically non-oxidized. The latter kind 
only is suitable for the ferric fluosilicate-hydro fluoric acid 
process. The analyses are given in Table 121. The last anal- 
ysis given for Lot 2 is the most accurate. 

TABLE 121. 





Lot 1. 


Lot. : 


>. 


Moisture 

Antimony on Hrv rp>idup 


20 % 
33.75% 

1.45% 
12.61 
12.08% 

160% 


47 . 75% 

39 . 22% 

2.25 

14.10% 
16.24% 




Copper on dry residue 


2.45% 
16. » 
17.20 


Arsenic 

Silver 


Bismuth. . . 


2.60 


Tellurium 




1 . 30% 

trace 


Selenium 







Iron 


0.50% 
12.06% 
16.00% 


' 9.8S%" 




Lead 

Fusol 


11.9 % 



The amount of ferric iron required for a given slime can be 
calculated from its composition, if the slime is unoxidized. or 
determined experimentally. The method of testing the iron- 
reducing power consists in boiling with an excess of ferric- 
sulphate and boiling the filtrate with metallic copper until 
all ferric iron is reduced. Multiplying the amount of copper 
dissolved by 1.76 and subtracting the result from the amount 
of ferric iron used gives the iron reduction figure. Of Lot 2. 
100 grams as dry slinie reduced 94.5 grams of ferric iron, which 
shows practically no air oxidation to have taken place. Of 
Lot 1, 100 grams as dry slime reduces 13. S grams ferric iron, 
showing approximately 85% air oxidation. This is about the 
usual figure for soft slime, dried in air. 

A test on 100 grams Lot 1 with ferric nuosilicate and hydro- 



APPENDIX. 



359 



fluoric acid gave a 25-gram residue, containing Fe 3.5%, Cu 
none, Sb 18.5%, Bi 3.45%, Pb 9.35%. 



TABLE 122. 





In Original Slime 


In Residue. 


Far Ceat. 

Dissolved. 


Antimony 

Copper 


27 grams 

1.16 grams 
10.1 
10.1 

1.28 


4 . 62 grams 
none 

2.34 
0.86 


82.9% 
100.0 


Arsenic 

Lead 

Bismuth 


76.5 
37.8 



Six hundred and fifty grams of Lot 1 (520 grams dry weight), 
leached with a fluosilicate-fluoride solution containing 42 grams 
ferric iron gave a 159-gram residue containing 39.8% silver, 
10.5% antimony, no copper, 24% lead, no arsenic, 3.41% 
bismuth. This shows an extraction of 90% of the antimony; 
all copper and arsenic; 42% of the lead, and 40% of the bis- 
muth. 

The poor extraction of lead was due to the solution con- 
taining too much HF, so that lead fluoride was formed and 
remained undissolved. The percentage of bismuth extracted 
is not of great importance, as the process recovers both un- 
dissolved and dissolved bismuth. The solubility of bismuth 
in these solutions was approximately 1 gram per liter. 

No further preliminary tests were thought necessary on 
Lot 2. 

The apparatus to be used consisted of a series of tanks. 
In the first the solution from the slime is electrolyzed with a 
low current density of about five amperes per square foot, 
using copper cathodes and antimony anodes. The antimony 
dissolves at the anodes while copper and presumably bismuth 
deposit. The solution is supposed to flow from the tanks 



360 



LEAD REFINING BY ELECTROLYSIS. 



practically free from copper and bismuth. The next series 
of tanks was much larger and contained lead anodes and copper 
cathodes; lead dissolving and antimony depositing, with a 
current density of about 12 amperes per square foot, which 
was found later to be decidedly too high, so more tanks were 
used and the current density reduced to 7 amperes. Leaving 
these tanks, the solution containing a little antimony passes 
through another somewhat smaller set, having lead anodes 
and cathodes. In the first of this set lead and antimony with 
some arsenic, and later lead and arsenic, and finally nearly 
pure lead deposit, or at least were expected to. The anodes 
in the antimony-depositing tanks contained about 0.6% anti- 
mony, and those in the arsenic-depositing tanks were of prac- 
tically pure lead. The dimensions of these tanks are given 
in Table 123. 



TABLE 123. 





No. 


Length, 
Inches. 


Depth, 
Inches. 


Breadth, 
Inches. 


Cathodes, 
Inches. 


Anodes, 
Inches. 


Copper tanks 

Antimony tanks .... 
Arsenic tanks 


6 
3 
3 


6 
11 
10 


8 
17 
16 


7 
14 
12 


6X6 

mxie 

10*X12* 


6X 7 
10X13* 
10X13* 



All tanks were fitted with independent agitators capable 
of maintaining a good circulation, which is very necessary with 
this process, because the solutions, except in the ferric iron 
tank, are very dilute with respect to the metals being deposited. 

The diaphragms for the ferric-iron tank were prepared by 
stamping a charcoal and asphaltum mixture into buttons, 
1J inches in diameter, and about A-inch thick, drying and 
baking below a red heat. About 2150 of these were inserted 
and made fast with thick asphaltum varnish in holes bored in 



APPENDIX. 361 

the sides of five wooden boxes, which were to form the anolytc 
compartments. These boxes were made of f-inch wood, and 
were 3 inches wide by 30J inches long by 22 inches deep inside. 
Before inserting the carbon buttons they were boiled in sodium 
nitrate solution to drive out the air and wet the buttons, so 
that they would finally become wetted through when elec- 
trolyte was added to the tank. The buttons also expanded a 
little by this treatment. The space occupied by the buttons 
on each side of each box was about 21J inches by 29J inches. 
As 215 buttons had an area of 264 square inches, the current 
density in the buttons averages about 2.5 times higher than 
the anode and cathode current density and approximated 
20 amperes per square foot. 

Five of the anolyte boxes were spaced with distance frames, 
about 3i inches apart in the clear between the boxes, in an 
asphalted wooden tank with an internal length of 42 inches, 
width 35 inches, and depth 24 inches. The whole was driven 
tightly together with wedges inside the tanks at one end, while 
the tank was securely braced outside to prevent its being 
strained by the pressure developed by the wedges. 

The anodes consisted of five sets of Acheson graphite rods, 
one inch in diameter and 24 inches long, cast in lead at the 
top and carried by reciprocating beams at the sides of the tank. 
The total motion was f inch. There were 19 anocles to the 
frame, spaced with 1J inches centres. The actual anode sur- 
face was about 4% greater than would be presented by a plane 
of the same overall measurements. The total anode area 
exposed was approximately 41.5 square feet. The six cathodes 
were of sheet lead 21X28 inches with an exposed total area 
of about 41 square feet. With a current of 330 amperes, this 
corresponds to a current density of about 8 amperes per square 



362 LEAD REFINING BY ELECTROLYSIS. 

foot. Provision was made to keep all the catholyte and 
anolyte in good circulation through the various respective 
compartments. The circulating apparatus adopted did not 
work well at all, unfortunately, with the result that the solu- 
tion in some of the anolyte boxes contained no remaining 
ferrous iron for a large part of the time, and the anodes after 
the runs were over were found to be considerably attacked 
in those places, although in other places, even. where as much 
current was used, there was no evidence of corrosion. 

The total cubic capacity of the electrolytic tanks, taking 
account of space occupied by electrodes and diaphragms, was 
about 25 cubic feet, while, when all tanks but the iron tank 
were cut out, the capacity approximated 15 cubic feet. 

The solution was made up originally by first dissolving scrap 
wrought iron in fluosilicic acid, and then treating 70 lbs. of 
oxidized slime of Lot 1 with a part of the solution. The solu- 
tions were then mixed together for the tanks and contained 
12.4 grams SiF 6 , 0.1 gram copper, 2.55 grams iron, 1 to 2 
grams HF, and 0.86 grams antimony per 100 c.c. 

The treatment of the slime of Lot 1 by SiF 6 and HF did 
not give as high an extraction as was expected from the tests 
made previously. A possible explanation is that the top of 
the barrel from which the slime was taken, differed in oxida- 
tion from the middle plane from which the sample was taken. 
A content of 2% or more of antimony was desired and had been 
expected. 

The total amount of the solution used was about 900 
liters. 

The slime treatment so far was not successful, but the 
solution was most easily prepared in that way, and that was 
really the reason this particular method was used. 



APPENDIX. 363 

The solution was fed first to the copper-depositing tanks, 
and the others were gradually brought into action as they 
filled up. 

The electrical conditions were about as follows, Table 124: 

TABLE 124. 

Average Volts. Average Current Density. 

Copper tanks, 0.17 5 amps, per sq. ft. 

Antimony tanks, .45, normally rose however to 2 volts 

at times, 7.4 " "."'.* 

Arsenic tanks, . 45, normally rose however to 2 volts 

at times, 1.7 " " " " 

The antimony anodes in the copper tanks dissolved regularly 
and evenly. The lead anodes in the antimony tanks dissolved 
without difficulty, but lead fluoride formed during the first 
run in patches on the surface, and collected as a white mud in 
the bottoms of the tanks. The lead anodes in the arsenic 
tanks of practically pure lead did not dissolve well at first. 
The surface was quite rapidly covered with an insulating layer 
containing lead fluoride. These were then replaced with 
anodes containing about 2% of antimony, with the idea that 
the anode slime of antimony would act as a diaphragm and 
keep the HF in the solution away from the anode surface. 
The anodes dissolved better thereafter. The explanation is, 
that the current is principally carried by the SiF 6 ion, the 
formation of lead fluoride being largely a secondary reaction 
between the PbSiF 6 formed, and the HF. 

The formation of lead fluoride in the tanks is not really 
necessary in the process, and did not occur afterward, but in 
the first part of the run the solution contained too much HF, 
and quite a little white lead had to be added to remove the 
excess. 



364 LEAD REFINING BY ELECTROLYSIS. 

The lead fluoride could be worked up by adding it to a 
batch of slime, when a reaction occurs as follows : 

2Sb + 3Fe 2 (SiFe 6 ) 3 + 3PbF 2 = 2SbF 3 + 6FeSiF 6 + 3PbSiF 6 . 

The copper tanks took altogether 6 to 16 amperes, arranged 
in two series, or 3 to 8 amperes per tank, with a current density 
of 3 to 8 amperes per square foot, and voltage of 0.2 to 0.24 
with 6 amperes per square foot. 

The antimony-depositing tanks took 60 to 150 amperes for 
the three tanks, or a current density of about 5 to 13 amperes 
per square foot, with normal voltage of 0.25 to 0.6. The de- 
posited metal was of various kinds, and no pure antimony was 
produced. The voltage rose much higher at times, and probably 
oxidized some antimony to the irreducible SbF . 

The "arsenic" tanks were operated with 10 to 100 amperes, 
averaging about 30, or a current density of 1 to 10 amperes 
per square foot. The voltage ranged from 0.5 to 1.5. 

The large ferric-iron producing-tank had an extremely high 
resistance at first, until the solution had penetrated the pores 
of the carbon. At the end of the run the temperature had 
risen to 36° C. and the current rose to 335 amperes, with 2.5 
volts. 

No difficulty was experienced with polarization at the 
anodes, provided they were kept moving back and forth by 
the mechanism provided therefor. Otherwise the tank would 
polarize in a minute or two, and the voltage would show an 
increase of from 0.6 to 0.8 volts. The counter electromotive 
force of the cell determined by opening the circuit and reading 
the voltmeter was about 1 volt. No difficulty with silica 
depositing on the anodes and causing polarization and gassing 
was experienced with this process as with the ferric-sulphate 



APPENDIX. 365 

process, and this could not very well happen, because the 
solution contained free hydrofluoric acid, which would, of 
course, keep silica in solution. 

Most trouble was caused by the carbon buttons loosening 
and dropping out in the tank. Part at least of this trouble 
was caused by faulty setting of the buttons. Some had been 
put in without any cementing material at all. The leaking 
holes were located and corked up, but still the efficiency was 
low, and at times the mixing of anolyte and catholyte was so 
rapid that the tank actually showed a loss of effect. By sam- 
pling different parts of the anolyte and titrating with per- 
manganate, the efficiency could be determined. The highest 
obtained for the whole tank was 56%, although three of the 
five anolyte boxes showed 100% at one time. 

The lead deposited at the cathode was of a peculiar char- 
acter. It was not solid, nor apparently crystalline, even under 
the microscope. It did not show any tendency to tree out, 
and make short circuits, but covered the cathodes in a felted 
layer, which would drop off when the layer became too thick, 
in say, twenty-four hours. The same kind of a lead deposit 
is that obtained from other solutions containing a fraction of 
a per cent of arsenic and antimony. As the lead was not in 
satisfactory shape to either build up a solid cathode or for 
melting, a rolling rig to pack the deposit down was made. 
This was not tried until the second run and then only for a 
time. 

The anodes were given about 50 complete vibrations per 
minute to keep them from polarizing. A good deal of the 
time there was no motion as the motor driving the anode 
frame was difficult to regulate with the means at hand. 

The difference in specific gravity of catholyte and anolyte 



366 LEAD REFINING BY ELECTROLYSIS. 

was only slight, during this run, but it seemed to increase as 
the percentage of lead in the anolyte diminished. Catholyte 
with 28 grams ferrous iron per liter, had a density of 1.132 
at 36°, while anolyte with 7.3 grams ferrous iron had a density 
of 1.144. 

Some slime of Lot 2 was treated during this run by anolyte 
taken from the ferric-iron tank. The slime was stirred up in 
a barrel with a slighter excess of ferric iron, calculated as 
follows : 

1 part copper requires 1.76 parts Fe' 
1 " antimony " 1.4 " 
1 " arsenic " 2.23 " 

1 " bismuth " 0.81 " 

1 " lead " 0.54 " ' 

The solution after settling was siphoned off and agitated 
with a small quantity of fresh slime to reduce any ferric iron 
or precipitate any silver in solution. The solution was then 
settled and run through a filter into a tub which fed the elec- 
trolytic tanks. The slime after treatment with the solution 
left only a small volume of a dense metallic residue, of far less 
bulk than the slime treated. It filtered fairly well with cold 
water, and washed rapidly with hot water. 

The residue was analyzed and found to contain lead 11.4%, 
antimony 14.1%, arsenic 1.48%, Bi 0.54%. The silver by 
solution in nitric acid and titration with NH 4 CNS was 58.3%, 
a little less than the actual amount. For quick determinations 
to control the process this method was used however. Taking 
silver in the original slime at 16.2%, obtained by the same 
method, and assuming that no silver was dissolved, the results 
are given in Table 125. 



APPENDIX 

TABLE 125. 



367 



In 13 Lbs. 
Wet Slime. 



Antimony 
Copper . . 
Arsenic . . 
Silver. . . . 
Bismuth . 
Lead . . . . 



2.67 lbs. 
0.15 lbs. 
0.96 lbs. 
1.11 lbs. 
0.18 lbs. 
0.68 lbs. 



In 1.9 Lbs. 
Dry Residue. 



0.27 lbs. 

none 
0.03 lbs. 
1.11 lbs. 
0.01 lbs. 
0.22 lbs. 



Extracted. 



90% 
100% 
97% 
none 
94% 
68% 



The solution from the slime treatment was partly passed 
into the series of electrolytic tanks, but mostly stored and 
used in the second run. 

The somewhat inferior results in extraction were probably 
due in part to the low temperature at which the slime treat- 
ment was conducted, namely, 12°-13° C. In the following 
run the temperature was 25°-30° C. 

A test was made on one-half barrelful of solution with the 
proper addition of slime, to see if there was any increase in 
the percentage of SiF 6 in the solution. It had been thought 
that the slime contained in an unrecoverable form products of 
the fluosilicic acid used in refining the lead. Very careful 
analyses before and after adding the slime showed no change 
in the amount of SiF 6 present. Very little could have been 
in the final residue, so that with well-washed slime there is 
no apprecaible quantity of fluosilicic acid or decomposition 
products left in the slime from lead refining. 

At the end of the run none of the tanks had given satis- 
faction. There was difficulty keeping the contacts in con- 
dition on the copper tanks, because the electrodes were very 
light. No pure copper was produced, and much pure antimony 
had been converted into impure metal. 

No good antimony had been made in the antimony tanks, 



368 LEAD REFINING BY ELECTROLYSIS. 

but the varying current density and composition and rate of 
feed of solution were so difficult to have controlled by my 
assistants before it was thoroughly understood what was re- 
quired, that anything different from a collection of all kinds of 
deposits on each electrode could not have been expected. 
The iron tank had failed because of internal leaks. 

The experimental plant was then shut down and altered 
in many respects. The contacts were improved, new and more 
powerful stirrers put in each tank, and the capacity of the 
antimony-depositing tanks increased 66%. 

The ferric-iron tank was taken apart, and the anolyte boxes 
tested by filling them with water, and all poorly set buttons 
taken out. Even after that fears were entertained lest the 
wood should expand or contract by wetting or drying and 
loosen the buttons. The plan of mounting the buttons in 
hard rubber plates by means of soft rubber rings cut from a 
rubber tube surrounding each button was considered, but it was 
thought to require too much time. To make sure of the success- 
ful operation of the tank, so that the process itself could be 
thoroughly tested, each anolyte box was covered with a double 
layer of cotton duck. The duck was so successful in with- 
standing the action of the solution, that it will undoubtedly 
itself provide a suitable and economical diaphragm if the 
tank is so constructed that new sheets of duck may be sub- 
stituted every month, say, and without its being necessary to 
take the tank itself apart. 

For the second run, the old solutions were analyzed before 
mixing, with results as in Table 126. 



APPENDIX. 
TABLE 126. 



300 



Fresh solution 

Old eatholyte 

Old anolyte after adding slime 

Old solution from antimony tanks .. . 
Old solution ready for depositing tanks 



Lead. 


Iron. 


SiF„. 


0.57 


2.14 


14.5 


1.13 


3.08 


12.8 


1.04 


3.03 


10.9 


3.18 


2.72 


12.1 


1.48 


2.96 


11.3 



Antimony. 



0.32 
1.46 
0.65 

0.88 



The mixed solution used contained about 2.85% Fe and 
12.6% SiF 6 and was maintained at about this strength through- 
out the run. The amount of HF present was not determined, 
but was not far from one per cent. The solution was entirely 
too weak for the best results, and was low in free acid, averaging 
about 1 or 2% only. What acid was not combined with ferrous 
iron was combined with lead, or the whole was combined with 
ferric iron. If the solution had contained more free acid, the 
power consumption on the iron tank would have been much 
less. It is rather surprising that such good results were ob- 
tained with such a weak solution. It had been intended to 
work with 16% SiF 6 , but one of the barrels in which acid had 
been stored had leaked out. 

In starting up, the first tanks to be put in operation were 
the copper-depositing tanks and one of the antimony -deposit- 
ing tanks. As the solution gradually filled the other tanks 
the current was increased. After twenty-four hours the iron 
tank at the end contained 4 inches of solution. As it filled 
the current was increased, keeping the voltage practically 
constant at 3.5 volts. The current reached 160 amperes after 
about 70 hours and the full 300 amperes was not reached for 
eight days. 

The tank could have taken the full current earlier, but 
was in series with the antimony and lead-arsenic depositing 



370 



LEAD REFINING BY ELECTROLYSIS. 



tanks, and the current was kept down in an attempt to get 
the desired pure antimony deposition. At that time the 
other tanks had been finally taken out, and thereafter the 
iron tank only was operated. 

The copper-depositing tanks did not give good results at 
any time, partly because the contacts were poor, and the cur- 
rent density on some electrodes was in consequence far too 
high. I scraped the deposits from two cathodes, one with a 
heavy deposit and the other with a light one. After melting 
they gave on analysis the figures in Table 127. 

TABLE 127. 



Lead. 



Copper. 



Bismuth. Antimony. 



Arsenic. 



Heavy deposit 
Light deposit . 



3.57% 
1.9 % 



4.15% 
10.5 % 



1-0% 
5-1% 



74 
69 



6% 
6% 



14.2% 
13.7% 



There were five antimony-depositing tanks on this run, 
instead of three as before. The highest current density used 
was about 6 amperes per square foot. Some of the best look- 
ing deposit contained 8.15% lead, so it was apparent that the 
most vigorous circulation and close regulation would be neces- 
sary for the successful production of antimony. 

The arsenic-lead depositing tanks gave nearly continuously 
a soft deposit of lead. 

After a few days running the copper tanks were found to 
be allowing copper to pass through them, and they were dis- 
connected and replaced by a box containing metal scraped 
from the cathodes in the antimony-depositing tanks. This 
box was 11 inches by 14 inches and the layer of metal which 
rested on a false bottom was about 3 inches thick. The solu- 
tion ran through too rapidly when the box was filled, and the 



APPENDIX. 



371 



ciippor was not all removed. From the analyses in Table 128, 
it will be seen that the lead in the box dissolved away first, 
precipitating antimony and copper, while later the antimony 
dissolved and copper precipitated. 



TABLE 128. 



Day of Run. 


Solution Fed to Copper 
Extractor. 


Solution Flowing from Copper 
Extractor. 




Cu 


Sb 


Pb 


Cu 


Sb 


Pb 


5th 

7th 

8th 


0.075% 

0.086% 


1.35% 
1.65% 
1.70% 
1.80% 


1.48% 
1.77% 


0.04% 
0.04% 


1.54% 
1-75% 
1.80% 


1.78% 


9th ... . 






0.044% 
0.01% 
0.012% 
0.017% 




11th 








12th .. . 










13th 





















Better results would have been obtained if the box had 
been filled with lead from the ferric -iron tank. This would 
have a finely divided form and be more active than the more 
solid metal that was used. With an arrangement such that 
the overflow of the box had been high enough to keep the pre- 
cipitating metal always flooded, and the flow of solution had 
been uniform, instead of intermittent, better results would have 
been obtained. 

Tests were then made to determine whether antimony and 
arsenic could be precipitated , by lead in a similar manner. 
The cathode lead in the ferric-iron tank, which was being 
packed down on the cathodes by rolling, was tried in the test. 
This material was found to analyze at two different times as 
follows : 



372 LEAD REFINING BY ELECTROLYSIS. 

TABLE 129. 

Sb 46% .30% 

As 47% .29% 

It consists of fine particles of lead loosely held together, 
with no crystallization apparent under a small microscope. 
It deposits in a felted layer on the cathodes and for use in 
precipitating was wiped from the cathodes by means of a 
trowel into a long tray resting on top of the tank. 

A layer of this lead about three-fourths of an inch thick 
was put in a funnel and solution from slime treatment rapidly 
run through, with results as given in Table 130. 

TABLE 130. 

Solution. Filtrate. Soft Material Left on Filter 

As 0.44% 0.18% 6.7% 

Sb 1.50% 0.87% 24% 

Pb 2.08% 6.9% 

The layer of precipitating lead was too thin and the speed 
of flow was rapid, so a complete extraction cf antimony and 
arsenic was not expected. The solution used contained some 
pentavalent antimony, which is not precipitable. 

After a number of other similar tests were made which 
showed a ready precipitation of arsenic and antimony by the 
cathode lead, the solution running from the copper extractor 
described above, already in use for two or three days, was 
passed through a 11-inch by 14-inch box containing a layer of 
cathode lead several inches thick, resting on a false bottom. 
The solution running through contained 0.07% As and 0.44% 
Sb. The antimony in the run-off came down with H2S only 
after a long time and with difficulty, showing that it was 
present in the solution as pentavalent antimony. 



APPENDIX. 373 

After some eight eon hours the solution running through 
began to contain more antimony, roughly determined by 
titrating 10 c.c. with permanganate solution, so another smaller 
box 7X7 inches with a layer of lead about 3 inches thick was 
put on just above the original box. All the electrolytic tanks 
except the main tank were emptied and their contents poured 
through with the solution coming from the slime treatment. 

The solutions running through the boxes were sampled every 
two hours for several days, and the samples analyzed for iron, 
antimony, and arsenic. It would take space unnecessarily 
to give all the results, but they showed a very good extraction 
of arsenic, and extraction of practically all the precipitable 
antimony, when sufficient lead was in the precipitating boxes. 

The solution still contained about 0.6% of antimony in 
the pentavalent condition, a result of either the high voltage 
developed at times in the antimony and lead-arsenic tanks, 
or of oxidation in some of the compartments of the ferric-iron 
tank, when the supply of ferrous iron was exhausted from 
insufficient circulation. I had another unfavorable condition 
to contend with in not having a sufficient stock of cathode 
lead on hand to fairly fill the precipitating boxes, as a result 
of which the lead in the boxes was at times practically ex- 
hausted before I had collected enough from the tank to fill 
them, while some was wasted as we were compelled to 
work. Practically the amount of lead taken from the elec- 
trolytic tank exceeds the amount dissolved from the precipi- 
tating boxes, because some lead is always coming into the 
system in the slime being added, but it is evident that it is 
necessary to have a certain stock on hand in the precipitating 
boxes, if all or nearly all of the antimony and arsenic is to be 
precipitated. What escapes, if any, has still to be electrolyzed 



374 



LEAD REFINING BY ELECTROLYSIS. 



near the cathodes of the electrolytic tanks, when it is largely 
removed in the cathode lead. 

The favorable condition of the cathode lead as a pre- 
cipitating material was due to the solution being somewhat 
impure in respect to antimony and arsenic, so with a complete 
extraction of these elements in the precipitating boxes, it would 
be necessary to run into the electrolytic tank a little solution 
still containing arsenic and antimony. 

The ferric-iron tank produced about 60 lbs. of granular 
lead daily, which was removed from the cathodes every 12 or 
18 hours and shoveled into precipitating boxes, of which there 
were eventually four in series. I sampled the four successive 
layers of one box which had been taken out, with the results 
given in Table 131. 

TABLE 131. 



Layer. 


Lead. 


Copper. 


Antimony. 


Arsenic. 


Top 

No.2 


none 
none 


present 


47% 
42% 
28% 
13.2% 


25% 
45% 
62% 
40.5% 


No. 3 . .. 




Bottom 













Of the four boxes in series, and while they were still operat- 
ing efficiently, I took samples as follows: Box A was 7x7 ins. 
and contained a layer about 4 ins. deep. Sample Ai was the 
top quarter, and A 2 , A 3 , and A 4 the following quarters. Box 
B was 11x14 ins. and contained a layer about 5J ins. deep. 
Samples B x to B 5 were of the five successive inch layers from 
top downwards. C was a pail 10J ins. diameter at the top 
and 8 ins. diameter at the bottom, and had a layer about 
4 ins. thick. Samples Ci to Cs are of the five successive layers 
from the top down. Box D was 11X9 ins. and contained a 



APPENDIX. 



375 



layer 9\ ins. deep. Samples Di to D 4 are from the four layers 
of equal depth. The analyses were hurried and the bismuth 
determinations were not satisfactory. In a general way, 
samples Ci and C 2 contained the most bismuth and a good 
deal of it, especially C x . C 3 is an accurate analysis by Mr. 
A. E. Knorr. The results are given in Table 132. 



TABLE 132. 



Number. 


Cu. 
Per Cent. 


Bi, 
Per Cent. 


Pb, 

Per Cent. 


Sb. 
Per Cent. 


As. 
Per Cent. 


Of Total 
Volume. 


A, 

A 2 

A 3 


40.2 
38.2 


none 
none 


trace 
none 
trace 
trace 
none 


35.7 

41.8 

52.5 

51.0 

52.5 ? 

64.5 

62.3 

68.2 

57.5 

38.5 

62.3 

59.5 
43.9 ? 
62.2 
56.6 


9.9 
12.3 
13.4 

16.7 

24.0 
25.1 
23.1 
25.7 

25.7 

24.0 

23.2 
14.6 
19.9 
13.7 
13.6 
11.1 ? 


2.3% 
2.3% 
2.3% 
2.3% 
8.0% 
8.0% 
8.0% 
8.0% 
8.0% 

2.4% 

2-3% 

2.2% 
2.1% 

2-1% 
9.9% 
9.9% 
9.9% 
9-9% 


A 4 

B x 

B, 


24.3 
2.5 


trace 
none 


B 3 [ none 

B 4 none 


none 


none 
none 
none 

} 2.0 

> none 

2.5 
2.0 

' 13% ' ' 
12 
13 
32.3 


B 5 

c, 

c 2 

C 3 

c 4 

c 5 

D x 

g 2 

D 4 


non e 

0.75 { 

none I 

none 
trace 
none 
none 
none 
none 
none 


none 
33% by 
difference 
13% by 
difference 

10.9 
2.2 
2.7 ? 

trace 

none 

none 

none 


64.5 
61.7 









There are evidently three distinct products, the first of 
which contains nearly all the copper, and would probably 
contain all the copper with a better arranged set of precipitat- 
ing tanks. This product, I believe, would be nearly pure 
copper and not a compound of copper with antimony or arsenic, 
as copper particles had been already formed, and it is probably 
only a question of time until the antimony and arsenic are 
all dissolved from the upper layers. This seems all the more 
probable since antimony and arsenic are known to precipitate 



376 LEAD REFINING BY ELECTROLYSIS. 

copper itself under the proper conditions. Further, the anti- 
mony and arsenic are in an ideal condition for chemical action 
on account of the fine division of the particles resulting origi- 
nally from their precipitation from solution, by lead particles. 
The complete absence of lead from this product and the follow- 
ing one is fortunate. Whether the copper product essayed 
40% or more, it would probably go to a copper anode furnace 
anyway. 

Bismuth is first found on going through the mass from the 
top downward, when the first layers containing lead are reached, 
and no other conclusion is possible, except that the bismuth in 
the solution run in is precipitated by lead and not by copper, 
antimony, or arsenic, and also that antimony and arsenic are 
precipitated by bismuth already precipitated itself by lead, 
while the bismuth redissolves and passes further clown until 
it comes into contact with metallic lead again. The bismuth 
must pass unprecipitated through the copper layers and the 
antimony-arsenic layers and be precipitated in the first lead. 
As this lead dissolves away in precipitating arsenic and anti- 
mony, the bismuth must dissolve with it only while the lead 
continues in solution and flows away, the bismuth is almost 
immediately again reprecipitated. Given a mass of precipitating 
lead into which the slime solution flows, the longer the time 
allowed, the wider will be the respective bands, and probably 
the higher will be the percentage of copper in the copper product 
and of bismuth in the bismuth product. Whether the bismuth 
layer would become in part at least pure bismuth is uncertain, 
but it makes very little difference as we have a simple and 
practical method of treatment. This is by stirring the product 
into the same solution as is used for treating slime, when the 
bismuth, being now more concentrated than in the slime. 



APPEND] 377 

will separate for the most part as insoluble bismuth fluoride, 
while antimony, arsenic, and lead dissolve, and the solution 
may be passed through the precipitating boxes with the slime 
solution. 

The principal product, in quantity at least, is the arsenic- 
antimony layer. This is fortunately free from lead. It shows 
no disposition to separate into two layers, one of antimony 
and the other of arsenic. This I proved by taking another 
sample, several days later, from the same locality that sample 
Bi was taken. It contains Sb 61%, As 26.7%. The pro- 
portion of the two metals is almost exactly the same as the 
proportion in which they are removed from the slime. 

At this time I have not yet had the opportunity of testing 
methods of recovering the antimony. If simply heated and 
melted in a crucible the product contains 17.2% As, and by 
further treating in a carbon crucible nearly to a white heat 
the arsenic is reduced to 8.7%. Exposing the melted metal 
to air does no good. 37.5 grams of alloy containing 8.7% 
arsenic was reduced by oxidation to 31 grams, but arsenic 
was still as high as 7.6%. Taking account of the vast difference 
in the boiling-points of metallic arsenic and antimony, and the 
absence of any strong combination between the two, sufficient 
heating of the antimony ought to give a complete separation. 
Possibly also by partial oxidation under the proper furnace 
conditions the arsenic can be got off as AS2O3 and the antimony 
left as metal. I expect to try heating the metal to the boiling- 
point of antimony in a carbon crucible placed in an electric 
furnace. Arsenic may be removed by treatment with sulphur. 

The various products can be distinguished by their appear- 
ance. The copper product is a black mud, the antimony- 
arsenic product is, when stirred with water in a glass vessel, 



378 LEAD REFINING BY ELECTROLYSIS. 

flaky like mica, and brilliant, while the bismuth product is 
intermediate between the antimony-arsenic product and the 
unchanged lead. 

The slime treatment itself was carried out in barrels, the 
method being to stir into the warm anolyte, containing about 
2.5% of ferric iron as it came from the electrolytic plant, a 
batch of slime that had been already used to reduce any excess 
of ferric iron left in the solution after treatment of the pre- 
vious batch. The slime and solution were stirred generally 
for about half an hour, using with our weak solution about 2\ 
cubic feet of solution for a 6-lb. lot of slime. After settling 
half an hour, the solution was decanted to another barrel with 
no silver or merely a trace in solution, and in one case when a 
test was made 0.16 gram of solid material per litre. To the 
decanted solution a fresh lot of slime was added to reduce any 
ferric iron. In this way each lot of slime and each lot of solu- 
tion was treated twice, so that the slime was thoroughly 
treated and the solution thoroughly reduced without its being 
required to get the exact quantities necessary for each treat- 
ment. If we were using too much slime the titration of the 
solution, after being put on slime for the first time, would rise 
with each batch, when the size of the slime lots could be di- 
minished. Some 45 lots were treated altogether, of which the 
first were removed separately while the last one-half or so 
were allowed to accumulate in the slime-treating barrel. The 
various lots were sampled and analyzed for silver by dissolving 
in nitric acid and titrating with NH 4 CNS solution. The 
results by this method, when checked up, were found a little 
low, say 1 to 3%. The figures are given in Table 133. 



APPENDIX. 



379 



TABLE 133. 



Lot. 


Per Cent 
Silver. 


Lot. 


PerCent 
Silver. 


Lot. 


Per Cent 
Silver. 


5 


56.4 
54.7 
40.0 
65.7 
61.3 
61.9 
57.8 
53.4 
61.6 
65.9 
63.1 


17 
18 
19 
20 
21 
22 
23 
24 
25 
26 
27 


62.9 
67.1 
62.9 
63.6 
59.6 
57.3 
60.3 
56.2 
61.8 
56.3 
61.8 


28 ] 

29 1 

30 f 

31 j 
281 
to I 

41 J 

42 1 
to I 

54 J 




6 




7 


67.2 


9 




10 




11 


67.4 


12 




13 

14 

15 and 16 

16 









The low percentage of silver in some few lots was due to 
the use of too much slime for a given amount of solution. An 
accurate analysis by Mr. A. E. Knorr gave for the original 
slime the figures given in Table 134. Our figures for Lot 28 
to 41 by his method and the percentages of extraction, on the 
assumption that the weights are inversely proportional to the 
percentages of silver, are also given. 

TABLE 134. 



Raw Slime. 



Treated Slime. 



Percentage of 
Extraction. 



Silver .... 
Bismuth . 
Copper . . . 
Lead .... 
Antimony- 
Arsenic . . 
Tellurium 



17.2% 

2.6% 

2.45% 

11.9% 

39.2% 

16.0% 

13% 



67.4% 
0.3% 
0.2% 
7.8% 

10.0% 
0.43% 
2.12% 



0.0 
97.0 
97.9 
83.1 
93.4 
99.3 

? 



Other analyses of treated slime are given in Table 135. 



380 LEAD REFINING BY ELECTROLYSIS. 

TABLE 135. 



Lot. 


Lead. 


Silver. 


Copper. 


Arsenic. 


Antimony. 


Bismuth. 


5 

6 

7 

9 

10 


8.3% 
8-5% 
15.6% 
5-2% 
7-4% 


56.4% 
54.7% 
40.0% 
65.7% 
61.3% 


none 
none 
none 
none 

none 


2-1% 
1-2% 

1-4% 
0.5% 

2-1% 


16.0% 
11-5% 
16.4% 
18.0% 
15.2% 


7.4% 
5-2% 
3.2% 
10% 
3.9% 



These early lots are not as representative of the process 
as the later combined lots. 

Regarding the percentage of extraction of the various 
metals there could be considerable variation even if there was 
complete oxidation by ferric iron. If the solution contained 
too much hydrofluoric acid the extraction of lead would be 
adversely affected,, as lead fluoride could separate in the slime. 
In this case also the percentage of bismuth extracted would be a 
minimum. On the other hand if the solution contained too 
little HF antimony would remain in the slime as trioxide in 
large quantities. In this case bismuth would be largely or 
entirely removed as fluosilicate. My results indicate that 
there is a safe mean between the two extremes. A ready 
method of control is to dissolve a sample of the treated slime 
in concentrated H 2 S0 4 . dilute to 500 c.c, add 50 cc. HC1, 
and titrate with permanganate for a rough antimony titration. 
If antimony is too high add a little more HF. 

From the standpoint of metal recovery, the extractions 
were pretty satisfactory, though it would be better if the 
silver residue was left in a purer condition. By the use of 
stronger and warmer solutions there would be an improve- 
ment to some extent at least, and longer agitation of slime 
with solution would probably help. Theoretically there is 
no reason why practically all the base metals could not be 



APPENDIX. 381 

removed, and possibly even tellurium could be removed and 
recovered by a certain procedure. 

The residue from the slime treatment is dense and solid 
and occupies a very small fraction of the space occupied by 
the raw slime itself. It filters rather slowly in the cold, but 
washes rapidly with hot water. 

Arsenic fumes were noted for a few hours during the first 
run, but no arsenic was evolved from the solution or apparatus 
after the copper, antimony, and lead-arsenic depositing tanks 
were cut out. 

For a commercial plant the following points are worth 
considering. 

The electrolytic tanks for depositing lead and producing 
ferric iron could use diaphragms of cotton duck stretched on 
wooden frames. The frames should surround the cathodes 
and not the anodes, as in my apparatus, because when renewals 
are required, which would probably be about once a month, 
the cathodes are more easily removed than the anodes. The 
frames or boxes would be pulled out, new duck stretched on 
and replaced. These boxes should be open at the bottom, and 
not quite reach the bottom of the tank. The heavy anolyte 
lying in the bottom of the tanks would dissolve any soft lead 
falling from the cathodes and prevent a troublesome accumu- 
lation, without affecting the cathodes, provided they did not 
reach too low in the tank. The anolyte with this construction 
could be readily circulated throughout the tank, a very de- 
sirable thing. The cathodes should best be of copper sheet 
and the tanks should be served by a crane so that the cathodes 
could be lifted out and away every twelve hours and the lead 
wiped off by the tank load, an operation that would take but 
a few minutes with apparatus like that shown on page 245. 



382 LEAD REFINING BY ELECTROLYSIS. 

The feed of solution would be divided as equally as possible 
between the cathode compartments, while the discharge would 
be merely through an overflow hose. A current density of 
10 amperes would probably be near the upper limit if it was 
desired to oxidize most of the ferrous iron. The voltage would 
be about two volts. 

The slime treatment probably need not be conducted in 
separate batches of regulated size. A whole day's production 
of slime could probably be placed in one tank, and anolyte 
from the tank allowed to collect there tor perhaps an 
hour, when it could be stirred and settled and the solution 
passed to the precipitating boxes. In this way the lead and 
bismuth might be removed first and the copper last, but as 
the precipitation of the metals is automatic there is no neces- 
sity for a constant composition of solution passing through the 
precipitators. 

The precipitation boxes should all have downward per- 
colation because the outflowing solution, containing more 
lead, is heavier than the inflowing. A little consideration shows 
that the metals should stratify horizontally under these con- 
ditions. By skilful regulation it is probable that the different 
products could be collected in separate boxes if desired. If 
not the different layers could be detected by the different 
appearance. The washing ought to be easy, if water is added 
at the top to displace the heavier solution, for the material is 
of a very open, pervious nature. The material in the boxes 
would be kept flooded at all times except when unloading 
by having the discharge at a level about the same as that of 
the top of the material. The resistance of the metal to the 
flow of the solution is very slight and is hardly to be con- 
sidered. 



APPENDIX. 383 

The solution flowed through one of the boxes in my experi- 
ment at the rate of 40 inches per hour, which was far too high. 
A speed of 4 inches per hour would not make the size of the 
precipitating tanks inconveniently large at all, and would give 
a better chance for the reactions to occur at the proper place, 
and would not allow the different metals to get beyond their 
respective zones of precipitation. 

The control of the process would be by titrating samples 
of the solution flowing from the electrolytic tanks by standard 
permanganate solution. The action on the slime can be followed 
in the same way. 

For following the operation of the precipitating tanks, 
titrating the inflowing and outflowing solution by permanga- 
nate, or taking the specific gravity of each, should give the 
desired information. In these titrations the permanganate 
oxidizes antimony and iron both, and will certainly oxidize 
arsenic in presence of HC1, and probably in presence of H 2 S0 4 . 

If the slime-treating solution accumulates lead fluosilicate, 
on account of the use of slime not thoroughly washed, an 
electrolytic method exists for taking this out again in the form 
of pure lead fluosilicate. 



INDEX. 

Acid fluosilicic, 30 

fluosilicic preparation, 174, 305 

hydrofluoric preparation, 174 

loss in evaporation, 323 

loss in refining lead, 32, 33, 35, 41, 42, 185, 253, 329 

loss in slime, 35, 270, 367 

loss on anode scrap, 270 

loss on cathodes, 38, 40, 269 

on surfaces, 37 
Alloys in anodes, 6, 54, 55 
Analyses, anode slime, 13, 57, 61, 99, 100, 116, 121, 133, 288, 289, 358, 359 

anode slime, treated, 379, 380 

dore bullion, 113, 160 

dore bullion from copper slime, 101 

dross from melting cathodes, 198, 202 

electrolytic antimony, 60 

material precipitated from slime solution, 374, 375 

refined lead, 13, 57, 284, 290, 298 
Analysis, methods of, 295 
Anode molds, 199, 202, 203, 316, 317 

molds, closed, 209, 317 
Anodes, casting, 203 

insoluble, of carbon, 361 

insoluble, of lead, for antimony-depositing, 144, 259 

scrap from, 255, 316 

storage of, 251 

sulphur in, 46 

tin in, 46, 47 

weight of, 316 
Anode slime, amalgamation of, 62 

amount of lead in, 53, 54 

bismuth in, 54, 76 

chlorination of, 68, 69, 71 

chlorination of alloys from, 67, 68 

drying, 256, 346 

385 



386 LEAD REFINING BY ELECTROLYSIS. 

Anode slime, extraction of metals from, 76, 79, 98, 100, 112, 113, 118, 130, 
134, 135, 137, 323, 356, 367, 379, 380 

from copper refining, 95, 100, 101 

fusion of, 71-73 

fusion of, products, 76, 77 

fusion of slags, 72 

fusion to alloy, 63-65 

iron-reducing power of, 358 

melting, see also Anode slime fusion, 63*-65, 71-79, 256, 257, 325 

melting with sulphur, 78, 79 

metals in, 46 

oxidation of, 96, 126, 128, 358 

physical condition of, 181 

polarized condition of, 49, 53 

porosity of, 48 

roasting, 128 

roasting with H 2 S0 4 , 129 

silica in, 35, 36 

treatment with combined fluosilicate and fluoride solution, 120, 121, 
125, 134, 355 

treatment with combined sulphate and fluoride solution, 116, 117 

treatment with copper fluosilicate, 125 

treatment with ferric fluosilicate, 355 

treatment with ferric salts of monobasic acids, 119 

treatment with fluosilicate solutions, 91 

treatment with sodium sulphide, 100, 123, 124, 323 

used as anode, 83, 85, 88, 126 

washing, 249, 271, 321, 322, 346 
Antimony, as precipitant for copper, 144, 145, 370, 371 

electrolytic refining of, 138 

extraction by ferric-fluosilicate + HF process, 379 

extraction by ferric-sulphate process, 112 

extraction by HF, 97 

fluoride electrolyte, 88, 138, 139 

in anode slime, 54, 56 

in melting slag, volatilization of, 74 

precipitation by lead, 371, 377, 382 
Antimony-depositing, anodes for, see Anodes insoluble. 

cathodes for, 259 

efficiency, 143, 324 

from fluoride solution, 135, 143 

from sulphide solution, 323 

tanks for, 259 

with soluble lead anodes, 360, 363, 367, 370, 379 
Arsenic in anode slime, 52 

in deposited antimony, 146 



INDEX. 387 



Arsenic in sulphate solutions, 102 
lead alloy, deposition of, 361 
precipitation by lead, 371 377, 382 

Arsenious acid in sulphate solutions, 102 

Ashcroft, E. A., S, 70 

Balbach, E., 15,"), 156, 158 

Benzenesulphonie acid, see Lead benzenesulphonate. 

Bibliography, 309 

Bismuth chloride electrolyte, 89 
in anode slime, 54, 76 
methyl sulphate electroylte, 89 
precipitation by lead, 375, 376 
solubility in fluoride solutions, 112 
solubility in sulphate solutions, 115 
recovery by ferric fluosilicate + HF process, 375 
recovery from sulphate solutions, 115 

Body, 102 

Borchers, W., 8, 124, 309 

Brewer, A. K., 1G0 

Bus-bars, 315, 345 

By-products from smelting, 291 

Cadmium-fluosilicate solution, 18 
Carhart, Willard and Henderson, 167 
Cascade system of tanks, 223, 241 
Casting anodes, 203, 212 

anodes in closed molds, 209 

cathodes, 230, 314, 319, 344 

lead from cathodes, 320, 346 
Catalysis of methyl acetate test, 19 
Cathode deposit, weight of per square foot, 40, 184 
Cathodes, casting, 230, 319, 344 

cleaning, 268-269 

for antimony-depositing, 259, 328 

for lead-depositing, 228, 229 

hanging, 319 

loss of acid on, 38, 40, 269 

melting, 320 

of deposited lead, 229 

placing in tanks, 319 

steel for lead depositing, 229, 267 

supporting bars for, 231, 233 
Chlorides, reduction by lead, 70 
Chlorination of alloys from slime, 67, 78 

of dry slime, 68, 69 



388 LEAD REFINING BY ELECTROLYSIS. 

Chlorination of wet slime, 71 

Chlorine storage, 70 

Cia Minera Fundidora y Afinadora, Monterey, Mex., 160 

Circulation of anolyte and catholyte ferric-iron tanks, 263 

of lead-depositing solution, 211, 237, 239, 344 
Cleaning anodes and cathodes, 268, 269 

tanks, 234, 268, 321 
Composition of lead-refining electrolyte, 41, 43, 187, 328, 346 
Condensers for hydrofluoric acid, 176 
Conductivity of various solutions, 17, 28 

determinations, 306 
Consolidated Mining and Smelting Co. of Canada, plant at Trail, B. C, 255, 

284, 312 
Contamination of cathodes by slime, 235, 285 
Contacts, 237, 315 
Copper addition to alloy from slime, 90 

anode slime, 95, 100, 101 

deposition of from slime solution, 131 

deposition with antimony anode, 360, 363, 367, 370 

fluoride electrolyte, 90 

in anode slime, 53 

lead alloy, treatment of, 291 

matte, 95 

matte leaching, 114 

matte roasting, 114 

precipitation by antimony, 144, 145, 370, 371 

process of Siemens and Halske, 93, 102 

scale, 114 
Cost of concrete tank, 219 

of depositing antimony, 148 

of glue or gelatine, 183 

of labor in tank-room, 272, 273 

of lead-depositing electrolyte, 242 

of lead-refining plant, 190, 191, 277, 283 

of making cathodes, 271, 319 

of making hydrofluoric acid, 177 

of melting lead, 273 

of molding anodes, 316 

of power influenced by current density, 187, 188 

of power influenced by solution composition, 188, 189 

of power lost in bus-bars, 227 

of refining lead, 272, 273 

of refining lead, comparative, 274-276, 279 

of steel cathodes, 229 

of treating slime, 191-196 

of unloading lead, 345 



INDEX. 389 



Cranes, 250, 313, 344 

Current density, consideration of, 1S3-191 
density, limiting, 53 

efficiency, lead-depositing, 209, 329 

Decomposition of fluosilicic acid by lead bases, 30, 246 

of fluosilicic acid by electrolysis, 33, 34 
Depreciation of tanks, 185 
Determination of conductivity of solutions, 306 
Diaphragms, 109, 110, 152, 262, 264, 265 

of carbon, 357, 360, 368 

of cotton, 368, 381 
Dietzel, Dr., 150-152 
Distillation of anode slime, 60, 62, 63 
Distribution of metals, extracting with H 2 SiF 6 + HF, 137 

ferric-sulphate process, 97, 98 

ferric-sulphate and HF process, 137 

melting slime, 76, 79 

roasting with H 2 S0 4 process, 133 
Dithionic acid, see Lead dithionate. 
Dore bullion refining, 149 

in furnace, 149, 150 
Drawing cathodes, 201, 285, 320 
Dross from melting cathodes, 198 
Drying slime, 256, 325, 346 

Easterbrooks, F. D., 155 

Efficiency of electric current affected by gelatine, 16 

of electric current in antimony deposition, 324 

of electric current in lead deposition, 329 
Electric furnace, 74, 75 

Electrolysis for ferric-sulphate solution, 101-109 
Electrolyte, introducing lead into, 243 
Electrolytic conductivity, 19 

refining rule of, 4 
Electromotive forces of solution of metals, 450-452 

forces of solution of alloys, 6 
Ethyl sulphuric acid, see Lead ethyl sulphate. 
Eurich, E. F., 274 
Evaporation, acid loss in, see Acid loss. 

lead-depositing electrolyte, 252, 323 

of fluosilicic acid, 29 

of water from electrolyte, 254 
Experimental tanks, 305 
Extraction of metals from slime, see Anode slime. 



390 LEAD REFINING BY ELECTROLYSIS. 

Factors for calculating ferric iron, 96 

influencing amount of lead in slime, 53 
Faraday's law, 3 
Ferric chloride for treating slime, 92 

fluosilicate for treating slime, 348 

sulphate, action of, 94 

sulphate for treating slime, 93, 94 
Ferrous sulphate, oxidation by air, 127 
Filtration of sulphate slime solution, 97 

of slime, 322, 346 
Fluoboric acid, see Lead fluoborate. 
Fluoride of antimony, etc., see Antimony fluoride, etc, 
Fluosilicic acid in hydrofluoric acid, 140, 147 

preparation, 178, 305 

see also Lead fluosilicate. 
Fluxes for melting to dore, 113 
Floors under tanks, 236 
Foundations for tanks, 233, 344 
Free HF in lead-refining solution, 30, 32, 36 

Gelatine, 14, 15, 22, 130, 158, 159, 170, 183 

quantity required, 42 

in silver-depositing electrolyte, 170 
Glue, see Gelatine. 

Haber, F., 309. 
Hanging cathodes, 319 
Hydrofluoric acid manufacture, 174 

yield, 243 
Hofman, H. O., 310 
Hofmann, O., 258 

Hook for lifting cathodes, Plate II, 337 
Howard skimmer, 198 

Inspecting tanks, 270, 326 

Interest charges* on lead, 183, 184, 276 

on silver and gold, 156, 157, 160, 170 
Inversion of cane-sugar test, 19 
Iron cathodes, 12 

in anodes, 46 

see also Cathodes, steel . 

Jacobs, E., 319 

Keith, N. S., 10 
Keith process, 10, 309 



INDEX. 391 

Kern, E. F., 27. 50, 54, 291, 293, 310, 311 

Labor for loading and unloading load, 272, 327 
for ma-king cathodes, 271, 319 
for melting lead, 273, 320, 827 

for tank-room. 272, 326 
for treating slime, 328 
Leaching apparatus, 257 

Lead acetate, refining solution, 10, 11 

alkaline solutions of, 11 

benzenesulphonate solution, 17, 19, 20-22 

chloride electrolyte, 7, 65. 69 

deposits, smoothness of, 16 

deposits, specific gravity of, 13 

dithionate preparation, 26 

dithionate solution, 17, 20, 22, 25 

ethylsulphate solution, 17, 19, 23 

fluoborate solution, 17, 20-22, 28 

fluoride, 8 

fluosilicate, conductivity, 28, 43, *5 

fluosilicate, crystallization, 31 

fluosilicate, preparation, 30 

fluosilicate solution, 17, 20-22, 32 

hydroxide theory, 12 

melting, see Melting. 

oxychloride and chloride bath, 8 

peroxide, 27, 58, 92, 93, 119-122, 154 

phenolsulphonate solution, 23, 24 

preparation of pure, 58, 59 

siphon, 197 

sulphide and chloride, 7 
Ledoux and Co., 301 
Levels in refinery, 197 
Limiting current density, 53 

Locke, Blackett and Co., Ltd., plant at Newcastle-on-Tyne, 183 
Losses at contacts, see Contact loss. 
Loss of acid, see Acid loss. 

Mattes of silver and copper, treatment of, 78 

from anode slime, 76-78, 79 
McNab, Alexander, 323 
Mechanical casting of lead, 202 
Melting lead, 198, 201, 202, 345, 320 

see also Anode slime. 

slime, 325 
Mennicke, H., 18, 309 



392 LEAD REFINING BY ELECTROLYSIS. 

Metals in slime, 46 

in solution, 46 
Method of analysis of: 

antimony, 297 

antimony fluoride electrolyte, 304 

dore bullion, 297 

electrolyte, 302 

hydrofluoric acid, 177 

matte from melting slime, 303 

refined lead, 298 

silica in slime, 304 

slags from melting slime, 302 

slime, 295 
Methyl sulphuric acid, preparation of, 168, 169 
Miller, J. F., 231, 311, 314 
Moebius, B., 155, 157, 158 
Molds for anodes, 202, 203, 316, 317 
Moving insoluble anodes, 103. 261 
Multiple system, 180-182 

Nebel, Moebius and, process, 159-164 

Operation of refinery, 167 

Ostwald, W., quoted, 18 

Oxidation of antimony at insoluble anodes, 135, 141, 142 

of electrolyte by air, 96, 239, 241, 243 

of ferrous sulphate by air, 127 

of slime by air, 126-129, 134, 256 

Patents, 309 

Philadelphia mint, silver refining, 159 
Platinum in anode slime, 54 
Polarization in lead refining, 189 

of anode slime, 49 

with insoluble anodes, 364 
Porosity of anode slime, 48 
Power cost, see Cost. 
Precipitation of fluosilicates, 141 
Pumps for lead, 316, 320 

for refinery solution, 239 
Purification of lead-refinery solution, 58 
Pyrogallol, 14, 15 

Refining dor6, see Dore. 

antimony, see Antimony. 



INDEX. 393 



Resorcin, 1 I 

Revolving cathode, 11, 19 

for fused bath, 9 
Rich lead, refining, 99 
Roasting slime, 12S, 129 

copper matte, 114 
Rosing lead pump, 197 
Rome, N. Y., plant at, 10 
Ryan, F. C, 246 

Saligenin, 14 

Sampling lead bullion, 213 

dore bullion, 298 
Scrap from anodes, 180, 244, 246, 255, 316 
Selby Smelting and Lead Company, 298 
Selenium, 94 

Senn, H., 17, 34, 36, 54, 294, 309 
Series system, 180, 182, 281-283 
Sherry, R. H., 8, 20 

Siemens and Halske copper process, 93, 102 
Silica in solution and slime, 32-36, 57, 98 

deposit on carbon anodes, 109 

use in slime fusion, 135, 257 
Silver amyl sulphate solution, 166 

deposition of solid, 165-167 

dissolved in roasting with sulphuric acid process, 130, 135 

distillation of, 63 

methyl sulphate solution, 152, 166 

perchlorate electrolyte, 167 

precipitation from ferric sulphate solution, 99 

sulphide converted to metallic, 78 
Slag from melting slime, 72, 76, 79 

reduction, 66, 82, 83 

treatment of, 80-83 
Slime, see Anode slime. 
Snowdon, R., 15 
Soda, use in slime melting, 71 
Sodium sulphide for treating slime, 100, 124, 323 
Specific gravity of lead deposits, 13, 16, 22 
Storage of anodes, 251 

of solution, 241 
Strength of acids, 18-21 
Sulphur in anodes, 46 

Tank systems, 180, 182, 227, 228, 241, 313, 343 
Tanks for antimony-depositing, 259 



394 LEAD REFINING BY ELECTROLYSIS. 

Tanks for lead-depositing, 213, 214 

for lead-depositing, size of, 214 

for lead-depositing of concrete, 215, 234, 235 

for lead-depositing of wood, 220 

for lead-depositing of wood, corrosion of bolts, 220, 224 

for making ferric sulphate, 260, 266, 306 

for slime treatment, 323 
Tellurium, 94 

Temperature of refining solutions, 44, 187 
Thum, F. A., 223 
Thum, Wm, 155, 171 
Tin nuosilicate solution, 18 

fluosilicate solution for refining, 47 

in anodes, 46-48 
Tommasi, D., 11 
Tray for catching drips, 319, 320 
Truswell, R., 209, 311 

United States Metal Refining Co.'s plant, 202, 255, 331 
Ulke, T, 310 

Valentine, W., 140, 231, 271 

Watt and Phillip, quoted, 10 
Washing cathodes, 39 

electrodes, 244, 345 

slime, 246, 271, 321, 322, 346 
Whitehead, R. L., 309 

Zinc chloride in fused-lead chloride, 8 



FEB 25 1908 



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